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IC 9137 



Bureau of Mines Information Circular/1987 



Eastern Coal Mine Geomechanics 

Proceedings: Bureau of Mines Technology Transfer 
Seminar, Pittsburgh, PA, November 19, 1986 



Compiled by Staff, Bureau of Mines 




UNITED STATES DEPARTMENT OF THE INTERIOR 



Information Circular 9137 



Eastern Coal Mine Geomechanics 

Proceedings: Bureau of Mines Technology Transfer 
Seminar, Pittsburgh, PA, November 19, 1986 



Compiled by Staff, Bureau of Mines 




UNITED STATES DEPARTMENT OF THE INTERIOR 

Donald Paul Hodel, Secretary 

BUREAU OF MINES 
Robert C. Horton, Director 





1 



Library of Congress Cataloging in Publication Data: 



Bureau of Mines Technology Transfer Seminars (1986 : Pittsburgh, 
Pa.) Eastern coal mine geomechanics. 

(Bureau of Mines information circular ; 9137) 

Bibliographies. 

Supt. of Docs, no.: I 28.27: 9137. 

1. Coal mines and mining- United States - Congresses. 2. Mine roof control -Congresses. 
3. Rock mechanics -Congresses. I. United States. Bureau of Mines. II. Title. III. Series: In- 
formation circular (United States. Bureau of Mines) ; 9137. 



TN295.U4 [TN805.A3] 622 s [622'.334] 86-607950 



PREFACE 

The papers contained in this Information Circular address the results 
of several research studies conducted by the Bureau of Mines in an ef- 
fort to improve mine roof support systems and mine design. Through 
eastern coal mine geomechanics research, new technology is being devel- 
oped that will enhance mine productivity, improve resource recovery, and 
increase mine workers' health and safety. 

The seven papers were presented at a technology transfer seminar on 
eastern coal mine geomechanics in November 1986. Technology transfer 
seminars are used by the Bureau of Mines to keep the minerals industry 
apprised of new technology and developments resulting from its research 
endeavors. Further information about current research efforts in other 
areas of concern may be obtained by writing the Bureau of Mines, Branch 
of Technology Transfer, 2401 E St., NW. , Washington, DC 20241. 



iii 



CONTENTS 

Page 

Preface i 

Abstract 1 

Introduction 2 

State-of-the-Art Testing and Analysis of Mine Roof Support Systems, 

by Thomas M. Barczak 3 

Characterization and Measurement of Longwall Rock Mass Movement, by Jeffrey M. 

Listak, John L. Hill III, and Joseph C. Zelanko 12 

Multiple-Seam Mining Problems in the Eastern United States, by Gregory J. 

Chekan, Rudy J. Matetic, and James A. Galek 27 

Integrated Design for Stability in Multiple-Seam Mining, by Chris Haycocks, 

Wei Wu, and Yingxin Zhou 44 

The Bureau of Mines Subsidence Research Program, by Michael A. Trevits, 

Roger L. King, and Bradley V. Johnson 57 

Subsidence Over Chain Pillars, by P. W. Jeran and V. Adamek 65 

Study of Dewatering Effects at an Underground Longwall Mine Site in the 

Pittsburgh Seam of the Northern Appalachian Coalfield, by Gregory E. Tieman 

and Henry W. Rauch 72 





UNIT 


OF MEASURE ABBREVIATIONS 


USED IN 


THESE PAPERS 


deg 




degree 


lb 


pound 


ft 




foot 


lb/cu ft 


pound per 
cubic foot 


ft 2 




square foot 


m 


meter 


f t/d 




foot per day 


min 


minute 


ft/ft 




foot per foot 


mm 


millimeter 


gal 




gallon 


ys/ft 


microsecond per foot 


gal/min 




gallon per minute 


pet 


percent 


gal/(min* 


acre) 


gallon per minute 










per acre 


psi 


pound per square inch 


gal/yr 




gallon per year 


psi/ft 


pound per square inch 
per foot 


gal/(acre 


•yr) 


gallon per acre 










per year 


psig 


pound per square inch, 
gauge 


h 




hour 


St 


short ton 


in 




inch 


yr 


year 


in/yr 




inch per year 







EASTERN COAL MINE GEOMECHANICS 



Proceedings: Bureau of Mines Technology Transfer 
Seminar, Pittsburgh, PA, November 19, 1986 



Compiled by Staff, Bureau of Mines 



ABSTRACT 

The Bureau of Mines is establishing design criteria for effective roof 
support systems and is developing technology to improve mine planning 
and design. Geologic studies, surface and underground mine monitoring, 
and laboratory evaluations were conducted. The seven papers in this 
proceedings Information Circular present results of several investiga- 
tions. Addressed are problems existing in both longwall and multiple- 
seam mines. Discussions include the effect of dewatering at a longwall 
mine site in the Pittsburgh coal seam and the effect of chain pillars 
on subsidence. Other results are measurement and characterization of 
longwall rock mass movement and an integrated design for stability in 
multiple-seam mining. 



INTRODUCTION 



The Bureau of Mines conducted a re- 
search study of coal mine geomechanics in 
the Eastern United States. The goal is 
to improve mine planning and design by 
establishing better roof and ground sup- 
port systems. Longwall mine roof fail- 
ures often occur because the support sys- 
tem provides inadequate ground control at 
the face. Better design, selection, and 
operation of such support systems may 
help decrease the capital risk involved 
in equipping a longwall face. Proper 
supports will also enhance productivity, 
maximize coal recovery, and minimize 
health and safety hazards. 

Problems in controlling ground are also 
prevalent in multiple-seam mining, which 
is practiced almost exclusively in the 
East. Past disregard for the sequence of 
mining multiple seams has led to major 
problems. Entry and pillar instability 
are caused by subsidence, pressure con- 
centrations, and stress zones. If mining 
proceeds without concern for adjacent 
coal seams, our coal resources will be- 
come depleted at a much faster rate. New 
technology will provide high recovery 
of multiple-seam coal resources in an 



economic and safe manner. To help alle- 
viate some of these problems, the Bureau 
conducted a series of studies. The ob- 
jective was to evaluate not only the ef- 
fects of the mining method, but also how 
geology relates to support requirements. 

A rock mechanics investigation moni- 
tored deformation of near-seam strata 
above a longwall panel in the Pittsburgh 
Coalbed. The study provided data con- 
cerning the caving mechanism associated 
with longwall extractions. Analysis of 
the data provided a better understanding 
of the interaction of strata behavior 
with longwall face supports. Another Bu- 
reau research study analyzed subsidence 
over raining operations. The purpose was 
to provide mine operators with a means of 
predicting movement of the surface ground 
and its effect on ground water. This re- 
search led to a computer model capable of 
predicting subsidence over longwall pan- 
els in the Northern Appalachian coal 
region. 

The report presents further details 
about the results of these research ef- 
forts and of other studies. 



STATE-OF-THE-ART TESTING AND ANALYSIS OF MINE ROOF SUPPORT SYSTEMS 

By Thomas M. Barczak^ 



ABSTRACT 



The Bureau of Mines is conducting re- 
search to evaluate the interaction of 
mine roof supports with the surrounding 
strata to provide for more effective de- 
sign and utilization of these structures. 
Scientific testing of a system requires 
consistent definition of pertinent param- 
eters and concepts. Various interpreta- 
tions of support capacity, as they exist 
in the mining community, are briefly dis- 
cussed in this paper, as are the defi- 
ciencies in historical uniaxial testing 
and analysis of mine roof support sys- 
tems. Critical load analyses show that 
structural failure of the support can 



occur at loads well below those of rated 
support capacity. The significance of 
support stiffness and post yield charac- 
teristics demonstrates the need to eval- 
uate supports and the strata as a system 
rather than independently. Efforts to 
develop supports as load-sensing devices 
and strata activity monitors are de- 
scribed. The benefits and limitations of 
finite element modeling as an alternative 
to physical testing are discussed. A de- 
scription of the Bureau's Mine Roof Simu- 
lator Facility and an overview of future 
research efforts are also provided. 



INTRODUCTION 



The development of more effective roof 
support systems and strata control strat- 
egies requires an accurate understanding 
of the interaction of roof support sys- 
tems with the geological environment in 
which they are employed. Improvements in 
roof support design and strata control 
strategies can be realized only if the 
response of the support system and the 
conditions under which it is loaded are 
understood. To achieve this understand- 
ing, the Bureau of Mines has constructed 
facilities to evaluate full-scale roof 
support systems under controlled loading 



conditions. Modern roof support systems, 
such as longwall sheilds, represent so- 
phisticated structures, which require 
more indepth analysis than the simple 
prop supports of the past. State-of-the- 
art testing and analyses of mine roof 
support systems at the Bureau of Mines 
include laboratory, field, and theo- 
retical studies using the most sophisti- 
cated equipment and analysis techniques 
available. This paper describes the Bu- 
reau's mine roof support facilities and 
research. 



SUPPORT CAPACITY AND WHAT IT MEANS 



One of the most fundamental measures of 
support performance is its load-bearing 
capability, which in the most elementary 
analysis is often ambiguously referred to 
as the capacity of the support. Several 
measures have been used to describe sup- 
port capacity, and it is necessary to un- 
derstand the differences and deviations 
of these measures in order to properly 



1 



Physicist, Pittsburgh Research Center, 



Bureau of Mines, Pittsburgh, PA. 



interpret the load-bearing capability of 
the support. 

One method of describing capacity 
for active supports, such as longwall 
shields, is the yield load of the hydrau- 
lic leg cylinders. In the case of shield 
sypports, the effectiveness of the legs 
to support a roof load is diminished by 
inclination of the leg cylinder from the 
normal to the plane of the canopy; only 
the vertical component of the leg force 
should be used to determine the capacity 



of the support. Capacity in this context 
is more precisely defined as roof-to- 
floor or vertical support capacity, being 
premised by two inherent assumptions: 
(1) The support is generally assumed to 
be uniaxially loaded with no face-to- 
waste (horizontal) or parallel-to-the- 
face (lateral) loading, and (2) vertical 
load application is assumed to occur at 
the location of the leg reaction. Kine- 
matically, there are components other 
than the leg cylinders that contribute to 
support capacity in structures such as 
longwall supports, and the actual capac- 
ity of the support should be determined 
from the total contribution of all compo- 
nents. Load applications at locations 
other than that of the leg reaction re- 
duce the capacity of the support by in- 
troducing moments. The presence of off- 
axis (horizontal or lateral) loading can 
either increase or decrease vertical sup- 
port capacity, depending upon the geom- 
etry and kinematics of the support 
structure. 

Another method used to describe support 
capacity is termed "yield load density," 
which is defined as the yield load of the 
support per unit area of supported roof. 
This measure should be considered only as 
an average of load density reacted by the 
support and not in the context of a uni- 
form pressure distribution of applied 
roof loading. It is computed simply by 
dividing the yield capacity by the area 
of supported roof. Since the yield ca- 
pacity is determined from a concen- 
trated load (or the equivalent stress 



distribution) acting at the resultant 
support reaction, assuming a uniform load 
distribution acting on the canopy surface 
of magnitude equal to the yield load den- 
sity will produce a support reaction in 
excess of the yield capacity. This means 
that support requirements should not be 
predicted simply from the weight of rock 
masses, assuming uniform load distribu- 
tion on the support elements or load con- 
centrations consistent with the center of 
gravity of the rock mass. The loading 
mechanism and interaction of the supports 
with the strata environment are discussed 
in a later section. 

Finally, it must be recognized that 
supports are rarely loaded uniaxially. 
While the primary function of the sup- 
port is to control roof-to-floor strata 
convergence, supports such as longwall 
shields are designed to control face-to- 
waste strata activity as well. Face-to- 
waste capacity of shield-type supports is 
generally rated at 30 pet of the rated 
vertical capacity, derived from an as- 
sumed 0.30 coefficient of the friction 
between steel and coal measure strata. 
The 30 pet of vertical capacity rating 
simply assumes that horizontal loads can- 
not be generated in excess of 30 pet of 
vertical loading since strata would then 
slip on the canopy interface. Actual ca- 
pacity of the support to resist horizon- 
tal loading may be substantially higher. 
Unlike vertical loading, which is re- 
lieved by hydraulic yielding, horizontal 
load capacity is determined by the struc- 
tural strength of suport components. 



CRITICAL LOAD ANALYSES 



As indicated in the previous section, 
support capacity is generally determined 
from load concentration at the canopy re- 
sultant reaction. Support capacity will 
be reached prematurely by application of 
loads of substantially smaller magnitude 
at locations other than the resultant 
support reaction. For example, while a 
500-st support could sustain a 500-st 
load concentrated at the leg (resultant) 
reaction, a load of less than 175 st ap- 
plied at the tip of the canopy would 



cause a typical two-legged shield support 
to yield. 

Hydraulic yield capacity is not the 
only measure of a support's load-bearing 
capability. Loading conditions exist 
that could structurally damage the sup- 
port with loads of magnitude well below 
the rated yield capacity. Critical load 
analyses are used to identify load con- 
ditions that cause maximum stress con- 
centrations in the support structure. 
Figure 1 illustrates symmetrical load 



Base 



Canopy 



Caving 
shiela 



1 I i i 



t t t t t t 

1 J 1 i 



k \\\Y\V 




I t t 

1 I 1 



Lemniscate 
links 





FIGURE 1.— Critical load test configurations. 

conditions that would cause maximum 
stresses in each major component of a 
generic, lemniscate shield. Several par- 
tial-contact conditions are illustrated 
in the figure by arrows, which represent 
load applications or restraint of the 
structure. Full-contact conditions are 



illustrated by a rectangle on the support 
member. Table 1 provides a description 
of the loading mechanism cases as well as 
a physical interpretation of the mine 
environment that could cause these load 
conditions. 

The results of these critical load 
tests for one particular shield, repre- 
sentative of two-legged shields in gen- 
eral, are described in the following 
paragraphs. 

CANOPY 

The canopy is the least stiff of all 
shield components and is therefore sub- 
ject to critical bending stresses at rel- 
atively small loads. Critical contact 
configurations suggest shield capacities 
of approximately one-half of rated hy- 
draulic yield capacity. In other words, 
loads of less than half of the rated hy- 
draulic yield capacity will cause stress 
concentrations approaching the yield 
strength of the steel under the contact 
conditions identified in table 1. 



TABLE 1. - Critical load tests 



Shield component 



Loading mechanism 



Physical interpretation 



Base. 



Canopy, 



Caving shield. 



Component link, 



Tension link. 



Create maximum bending stress in base 
by supporting base as a simple beam 
with canopy inducing concentrated 
force at center span of base. 

Same mechanism as for base with 
bending created in canopy. 



Simulate horizontal displacement of 
canopy relative to base by creating 
horizontal couple as shown in 
figure 1. 

Same strategy as with caving shield. 
Maximum loading occurs when canopy 
rotation is prevented, causing 
shield reactions at hinge pin and 
in caving shield. 

Same strategy as witb caving shield. 
Maximum tension link loading occurs 
when links are in horizontal con- 
figuration (low shield heights). 



Base supported on toe and 
rear with leg force 
causing bending of base 
structure. 

Same physical description 
as for base with canopy 
simply supported at tip 
and hinge. 

Canopy and base promoting 
horizontal reactions due 
to vertical shield 
convergence. 
Do. 



Do. 



BASE 

The base is the stiffest of all shield 
components, and no contact configurations 
were found to induce critical stresses 
for this particular shield. 

CAVING SHIELD 

Roof-to-floor convergence of the 
shield, which is thought to be the pre- 
dominant loading mechanism in an under- 
ground environment, produces relatively 
little loading in the caving shield. 
However, horizontal displacement of the 
canopy coupled with a horizontal re- 
straint of the base to cause relative mo- 
tion between the canopy and base cre- 
ates stresses approaching yield strength 
of the caving shield material with ver- 
tical loads below the hydraulic yield 
capacity of the support. The effect 
of gob loading directly on the caving 



shield 
study. 



was not investigated in this 



LEMNISCATE LINKS 



No contact condition was found to in- 
duce more than 6,000 psi of stress, which 
is less than 20 pet of the material yield 
strength. Apparently much of the energy 
is absorbed by the leg cylinders and in 
caving shield bending, with little en- 
ergy being transferred to the lemniscate 
links. The primary function of the links 
under these test conditions is simply 
as a guidance system to control canopy 
canopy motion. 

It should be noted that these test re- 
sults are for one shield only. Other 
shields may behave differently depending 
on their structural characteristics, but 
the critical load conditions should be 
applicable to most shields of this basic 
configuration. 



INTERACTION OF THE SUPPORT WITH THE STRATA 



In the evaluation of roof support sys- 
tems it is important to understand that 
the roof support elements and the sur- 
rounding strata act as a system, respond- 
ing to changes in the physical mine envi- 
ronment due to extraction of coal and 
associated redistribution of stresses. 
The strength (capacity) of a roof support 
element alone is not a meaningful measure 
of support performance. The loading of a 
roof support element can be significantly 
dependent upon the stiffness of the sup- 
port structure, as a stiffer structure 
will receive a higher load from converg- 
ing mine strata than will a softer struc- 
ture. This concept is illustrated in 
figure 2, where two hypothetical struc- 
tures of different stiffness, Kj and K.2 , 
with &i < K 2 , are displaced an equal 
amount. As shown, the resultant load on 
the stiffer structure would be 1.5 times 
that on the softer structure under the 
same load condition. 

The significance of support stiffness 
for behavior of the support is most cri- 
tical in passive roof support systems, 
such as posts and cribbing. Recently, 
concrete and steel-fiber-reinforced (SFR) 
concrete have been used as substitute 



materials for wood in crib applications. 
The higher compressive strength of con- 
crete compared to wood (about eight 
times) results in a much higher yield 
strength (capacity) for cribs constructed 
of concrete. However, as seen from the 
load-displacement characteristic of wood 
and concrete illustrated in figure 3, 
concrete produces a much stiffer crib 
structure than does wood. These results 
indicate that a solid SFR concrete crib 




DISPLACEMENT^) 
FIGURE 2.— Significance of support stiffness. 



1,200 



900 



600 



300 



-SFR concrete 



Nonreinforced 
concrete 




-Wood 



^--wooa 



4 6 8 10 

DISPLACEMENT, in 

FIGURE 3.— Crib loading characteristics. 



would experience a load 18 times that 
of a wood crib subjected to the same 
displacement prior to reaching yield 
strength of the crib. The displacement 
at which the yield strength of the sup- 
port is reached is also a critical mea- 
sure of support performance. The stiff 
concrete supports, despite having a much 
greater yield strength, reach yield load 
at much smaller displacements than cribs 
constructed of wood. Tests indicate con- 
crete cribs fail at less than 0.5 in of 
displacement, whereas wood can displace 
nearly 3 in before reaching yield load 
and over 10 in before failing. There- 
fore, if the convergence of the roof is 
irresistible (displacement loading), the 
high yield strength of concrete cribs 



will be of no advantage if the roof con- 
vergence exceeds 0.5 in. 

Equally important to the loading char- 
acteristics of passive roof support sys- 
tems is postyield behavior. As seen in 
figure 3, the stiff concrete supports 
lose load-carrying capability almost im- 
mediately after reaching yield load, 
whereas wood continues to provide support 
resistance for several inches of dis- 
placement. In terms of energy absorp- 
tion, the wood crib can absorb more en- 
ergy than a concrete crib owing to its 
ability to deform and maintain resistance 
to load. The idealized roof support 
should be initially stiff to be rapidly 
load bearing, should have yield strength 
sufficient to support the deadweight re- 
sponse of the immediate fractured strata, 
and should have sufficient postyield 
characteristics to be compatible with the 
displacement behavior of the overburden. 

The stiffness of an active roof sup- 
port, such as a longwall shield, is also 
critical to the interaction of the sup- 
port with the strata. A stiff structure 
that generates high reactive forces can 
cause unnecessary fracturing of incom- 
petent roof or floor strata or cause 
further instability of unstable strata. 
While the hydraulic support is self- 
relieving, the yield load of modern pow- 
ered supports ranges from 250 to 800 st; 
this is sufficient to cause pressure dis- 
tributions on the canopy and base that 
could damage incompetent roof or floor 
strata. 



SUPPORTS AS MONITORS 



Roof support elements, both passive and 
active structures, can be used as roof 
load monitors by proper instrumentation 
of the support structures. Past efforts 
to use roof supports as load-sensing 
devices have generally been limited to 
one-dimensional analysis by summation 
of measured forces to obtain a verti- 
cal (roof-to-floor) reaction. As previ- 
ously indicated, supports are usually not 
loaded uniaxially. Modern longwall sup- 
ports, such as the shield, are designed 
to resist both roof-to-floor (vertical) 
and face-to-waste (horizontal) loading, 



requiring two-dimensional analyses to 
determine vertical and horizontal load 
reactions. The Bureau has demonstrated 
that vertical and horizontal support re- 
sistance of a longwall shield can be rea- 
sonably determined from static rigid body 
analyses by measurement of leg, canopy 
capsule, and compression lemniscate link 
forces. The two-dimensional model as- 
sumes no gob loading, nor does it recog- 
nize any out-of-plane loading or the ef- 
fect of moment loading due to imbalances 
in the leg and lemniscate link forces. 
From controlled tests in the Bureau's 



Mine Roof Simulator, vertical shield 
loading could be predicted to within 3 
to 5 pet and horizontal loading to within 
25 pet. 2 

In theory, the limitations of the two- 
dimensional model can be overcome by 
three-dimensional modeling of the support 
structure to account for out-of-plane 
forces and moments. Unfortunately, the 
advancement of a three-dimensional model 
is difficult since the accumulation of 
unknowns far outpaces the available force 
and moment equilibrium equations, making 
the system statistically indeterminate. 
While elimination of unknowns with rea- 
sonable engineering judgments as to non- 
participating forces produces a solu- 
tion, the shield support is primarily 
designed to resist loading in two dimen- 
sions (roof-to-floor and face-to-waste), 
making the two-dimensional model adequate 
for most analyses. 

It is also recognized from tests with 
shield supports that vertical conver- 
gence produces both a vertical and a hor- 
izontal load reaction; likewise, a 
horizontal displacement produces both a 
horizontal and a vertical support reac- 
tion. In other words, the shield support 
not only reacts axial loads in direct 
response to strata displacements (shield 
convergence), but also induces offaxis 
load reactions as a result of the mechan- 
ics of the shield structure. Only load 
reactions resulting from strata activity 
need to be resisted for successful appli- 
cation of the longwall method; support- 
induced loads are indications of ineffi- 
cient designs if they are of no benefit 
to strata control. To put this in per- 
spective, in situ tests indicate that 
over 75 pet of horizontal load experi- 
enced by shield supports is the result of 
setting the support against the roof. 
There is also some evidence that a sig- 
nificant portion of the remaining 25 pet 



is due to vertical roof convergence and 
not face-to-waste strata activity. 

To resolve the issue of the source of 
horizontal shield loading and to develop 
methods whereby the supports can truly be 
used as monitors of strata activity, the 
Bureau is investigating the utilization 
of a linear elastic model with two de- 
grees of freedom. Mathematically, the 
model is expressed as follows: 



F h 



" K l 



+ K 2 
+ K, 



(1) 
(2) 



where 

F v = vertical support resultant load 
reaction, 

F h = horizontal support resultant load 
reaction, 

v = vertical shield displacement, 

h = horizontal shield displacement, 

and 

I , K 2 , K3, K4 = stiffness coefficients, 



K 



By controlled uniaxial displacement 
tests in the Bureau's Mine Roof Simula- 
tor, the stiffness coefficients are de- 
termined for a particular shield. Using 
numerical values for the stiffness param- 
eters, the inverse of equations 1 and 2 
can be used to determine shield displace- 
ments, which are indicative of mine roof 
convergence and strata activity, if re- 
sultant shield loading is accurately 
known. This also enables horizontal 
loading produced by face-to-waste strata 
activity to be distinguished from shield- 
induced horizontal loading due to verti- 
cal roof convergence. 



MATHEMATICAL MODELING— AN ALTERNATIVE TO PHYSICAL TESTING 



Mathematical models provide a simple, 
cost-effective method to evaluate load 
responses of structures without the need 

2 Barczak, T. M., and R. C. Garson. 
Technique To Measure Resultant Loading on 
Shield Supports. Paper in Rock Mechanics 



for full-scale physical testing. While 
simple rigid body statics may be used to 
determine the elementary behavior of a 

in Productivity and Protection (Proc. 
25th Symp. on Rock Mech.). Soc. Min. 
Eng.,1984, pp. 667-679. 



support in relation to an applied load, 
more sophisticated analyses are required 
to analyze the load-displacement rela- 
tionship of a support structure. The 
linear elastic model presented in the 
previous section is a first step in eval- 
uating roof supports as elastic bodies. 

One of the most powerful techniques 
used to analyze structures is finite ele- 
ment modeling. According to its basic 
principle, a structure is idealized as a 
composition of a number of finite pieces 
rather than continuous elements. This 
concept enables the step-by-step buildup 
of the load-displacement relationship of 
a structure as a whole from those basic 
elements of which the structure is com- 
posed. The accuracy of the model is de- 
pendent upon the number of elements as 
the stresses are averaged over the area 
of the elements. Proper selection of 
elements and boundary conditions is re- 
quired because different types of ele- 
ments behave differently. Knowing the 
geometry and elastic properties of the 
elements allows computation of structural 
deformations for each of the finite ele- 
ments from which areas of critical stress 
concentrations can be identified. 

The Bureau has used finite element 
analyses in the following ways in mine 
roof support research: 

Identification of load cases for test- 
ing . - The critical load tests presented 
in the section on critical load analyses 
were determined from a simple two-dimen- 
sional finite element model. 

Location of strain gauges for physi- 
cal testing . - It is generally desirable 
to locate strain-measuring instrumen- 
tation at areas of critical stress 
concentrations. 

Identification of critical stress con- 
centrations . - Three-dimensional models 
have been used to identify critical 
stress concentrations in major support 
components. 

Evaluation of canopy pressure distribu- 
tion from strain contours on the canopy 
structure . - Measurement of canopy pres- 
sure distribution by analysis of under- 
side strain contours proved difficult 



owing to multiple contact conditions that 
produce similar strain profiles and to 
difficulty in modeling partial contact 
conditions. 

Finite element modeling of supports has 
proved more challenging than originally 
anticipated. The general behavior of the 
structure in terms of load distribution 
can be determined from finite element 
modeling, but an accurate picture of load 
transfer in the shield structure is some- 
times difficult to obtain. Mathematical 
modeling requires an accurate definition 
of the problem, including both the prop- 
erties of the structure and the load and 
boundary conditions. Discrepancies in 
finite element model predictions are of- 
ten caused by improper modeling of load 
conditions. This is particularly diffi- 
cult in modeling support behavior because 
of the multitude of contact configura- 
tions that exist with rock strata inter- 
faces. Longwall supports are basically 
crude pin-jointed structures, and model- 
ing pin frictions in joints can be diffi- 
cult. Factors such as friction, ability 
to accurately model internal platework, 
and the effect of stress concentrations 
at discontinuities can have a significant 
impact on results. 

The Bureau's experience with finite 
element modeling of longwall shields is 
that simple beam models constructed to 
fit a known shield response from physi- 
cal testing are more successful and much 
cheaper than complex models, which may 
geometrically look more like the support 
structure, but are very difficult to con- 
struct to properly simulate the struc- 
tural fabrication of the shield. 

A further implication of the Bureau's 
research in finite element modeling of 
roof supports is the discovery that com- 
plicated shield component constructions 
are sensitive to some unique load appli- 
cations. While current supports are suf- 
ficiently overdesigned that these unique 
load conditions do not represent prob- 
lems, efforts to improve designs from a 
stress optimization viewpoint will re- 
quire more refined load analyses than the 
relatively simple criteria presently used 
in support design. 



10 



BUREAU OF MINES FACILITIES 



The heart of the Bureau's roof support 
test facilities is the Mine Roof Simula- 
tor (MRS) illustrated in figure 4. The 
MRS is unique in the world in its abil- 
ities to apply both a vertical and a hor- 
izontal load simultaneously. 

Both the vertical and horizontal axis 
can be independently programmed to oper- 
ate in either force or displacement con- 
trol. This capability permits tests such 
as true friction-free controlled loading 
of shields, which cannot be accomplished 
in uniaxial test machines since the 
shield reacts a horizontal load to ver- 
tical roof convergence. Friction-free 
tests of this nature can be accomplished 
in the MRS by allowing the platen to 
float in the horizontal axis by command- 
ing a zero horizontal load condition. 
Likewise, the MRS can apply controlled 
horizontal loading to a shield support, 
whereas uniaxial test machines can apply 
only vertical loading with no control 
over horizontal load reactions or capa- 
bility to provide a specified horizontal 
load to the structure. 

The machine incorporates 20-ft-square 
platens with a 16-ft vertical opening, 
enabling full-scale testing of longwall 
suport structures. Capacity of the simu- 
lator is 1,500 st of vertical force and 
800 st of horizontal force and controlled 
displacement ranges of 24 in vertically 
and 16 in horizontally. Load and dis- 
placement control is provided in four 
ranges operating under a 12-bit analog- 
to-digital closed-loop control network, 
providing a load control capability of 
better than 0.1 kip (100 lb) and dis- 
placement control capability of better 
than 0.001 kip in the smallest load- 
displacement range. 

Machine control and data acquisition 
are achieved by a DEC 11/34 computer. 
Eighty-eight channels of test article 




FIGURE 4.— Mine roof simulator. 

transducer conditioning are provided. 
Data acquisition is interfaced with the 
control network so that machine behavior 
can be controlled by response of the test 
article instrumentation. For example, 
tests can be terminated or held when 
strain values reach a designated level in 
specified areas of the support structure. 
High-speed data acquisition is available 
with a separate DEC 11/23 computer at a 
rate of 300 samples per second. An X-Y-Y 
recorder provides real-time plotting of 
three data channels, and all data are 
stored on computer disks for subsequent 
processing and analysis. 



FUTURE DEVELOPMENTS AND CONCLUSIONS 



The Bureau of Mines plans continued re- 
search on mine roof support systems in an 
effort to develop more cost effective and 
safer systems for the mining industry. 



Several aspects of ground control, sup- 
port design, and strata interaction are 
not yet well understood. The Bureau 
hopes to advance the state-of-the-art of 



11 



ground control and support testing with 
the following goals: 

1. Determine the effectiveness and 
limitations of finite element analysis in 
support design and behavior. Recommend 
guidelines for advanced analysis of mine 
roof support systems using finite element 
analysis. 

2. Continue efforts to develop tech- 
niques to effectively use supports as 
strata monitors to develop a more basic 
understanding of strata activity so that 
support requirements can be more closely 
engineered to strata behavior. 

3. Continue to explore and propose 
concepts of support behavior, such as the 
significance of support stiffness in load 
behavior, which may not be apparent to 
mine operators, support manufacturers, or 
other researchers. 



4. Continue to explore advanced tech- 
nology applications to current and novel 
mine roof support concepts or support 
designs. 

In conclusion, ground control is cri- 
tical to successful mining operations. 
Nearly all of the financial risk associ- 
ated with a mining venture can be attrib- 
uted to ground control. Current mining 
operations require capital-intensive sys- 
tems costing several million dollars for 
each installation, and roof support fa- 
talities continue to be the number one 
killer. Testing and analysis of roof 
support systems is necessary to provide 
better utilization of support systems and 
the development of improved designs for 
the future. 



11 



CHARACTERIZATION AND MEASUREMENT OF LONGWALL ROCK MASS MOVEMENT 
By Jeffrey M. Listak, 1 John L. Hill III, 1 and Joseph C. Zelanko 1 



ABSTRACT 



The Bureau of Mines has conducted a 
rock mechanics study to monitor deforma- 
tion of near-seam strata above a longwall 
panel in the Pittsburgh Coalbed. The 
primary goal of this research was to de- 
termine the height of caving immediately 
behind advancing longwall face supports. 
This study, although site specific, pro- 
vides information on the caving mechanism 
associated with longwall extractions so 
that strata behavior and its interaction 
with longwall face supports can be better 
understood. 

Two vertical boreholes, positioned 550 
ft apart along the centerline of a long- 
wall panel, were drilled from the surface 
to intercept the coalbed at a depth of 
approximately 650 ft. Various downhole 



geotechnical instruments were used to 
monitor strata deformation. In addition, 
surface elevation surveys were conducted 
to differentiate between surface and sub- 
surface activity. 

This report discusses the caving char- 
acteristics of the strata as the longwall 
panel approached and passed beneath the 
boreholes. Physical property data are 
also presented to demonstrate the rela- 
tionship between caving behavior and lo- 
cal geology. Data show that immediate 
caving of strata above the longwall face 
occurred at a height of less than 23.5 ft 
and that strata behavior above longwall 
extractions is highly dependent upon 
lithology, with major disturbances occur- 
ring at weak lithologic zones. 



INTRODUCTION 



For a safe and productive longwall op- 
eration, the optimization of both roof 
control and operational efficiency is es- 
sential. However, these two major con- 
tributors to longwall success are depen- 
dent upon the accurate prediction of roof 
support capacity requirements. Longwall 
failures of the past have been attributed 
to lack of understanding of roof and 
floor behavior and poor caving character- 
istics in the gob area. 

Historically, roof caving in the gob 
has been evaluated by several methods. 
One method is a theoretical approach 
based on the cantilever beam theory of 
structural mechanics. This approach as- 
sumes that the stratified roof rock lay- 
ers act as cantilever beams above the ex- 
tracted coal. By knowing the material 
properties and the thickness of the lay- 
ers, the extent of the interlayer loading 

'Mining engineer, Pittsburgh Research 
Center, Bureau of Mines, Pittsburgh, PA. 



can be determined (1_). 2 As the roof lay- 
ers fail in stepwise fashion, caved mate- 
rial swells or bulks to fill the void and 
the caving height can be calculated. 

Rock classification is another method 
used to characterize mine roof for cave- 
ability prediction. Much of this work 
has been done in the European coalfields 
by Pawlowicz, Kidybinski, and Kostyk of 
Poland and by Proyavkin and Davidyanc of 
the U.S.S.R. as cited by Kidybinski (2_). 
In addition, Ghose (3_) used geotechnical 
logs to classify mine roof for cave pre- 
diction in coal mines in India. Ghose 
also cited the notable classification 
systems proposed by Vasiliev et al. and 
Nenasheva et al. of the U.S.S.R. and 
the Roof Quality Index of Bilinski and 
Konopko of Poland. Bieniawski (4_) devel- 
oped an engineering classification of 

^Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this paper. 



13 



rock masses called the Geomechanics Clas- 
sification or rock mass rating (RMR). 
This classification system utilizes six 
parameters, which are measurable in the 
field or which can be obtained from bore- 
hole data, to determine the "stand-up 
time" for an unsupported roof span. 

More recently, numerical modeling has 
been used to provide good estimates for 
cave prediction (e.g. , the finite ele- 
ment method, distinct element method, and 
boundary element method). 

Although these methods are of value 
since they offer guidance, precise iden- 
tification of the caving behavior of the 
strata is still unresolved. Very few 
systematic studies that actually moni- 
tored the caving mechanism associated 
with longwall mining have been conducted. 
Field investigations (_5~J7) performed to 
directly measure bed separation over 
longwall extractions have had limited 
success. However, one such study con- 
ducted recently (_5) used a borehole cali- 
per tool to measure bed separation and 
had conclusive results that are very sim- 
ilar to the findings stated later in this 
report. 

To develop a better understanding of 
longwall strata behavior and its inter- 
action with longwall face supports, the 
Bureau of Mines installed various geo- 
technical instruments in two boreholes 
drilled over a longwall extraction lo- 
cated in southwestern Pennsylvania. This 
study is intended to lay the groundwork 
for additional research in order to de- 
velop a data base for strata behavior 
above longwall extractions. The informa- 
tion from these studies could effectively 
improve the method for selecting longwall 
roof support capacities. 



Several analytical methods (8-10) have 
been developed to predict longwall strata 
behavior and associated face support 
loading. These methods for selecting the 
proper load capacity of longwall roof 
supports are based on the assumption that 
some finite volume of roof material, of- 
ten assumed to be cubic or parallelpiped 
in geometry, is being held by the sup- 
port. The boundaries of this volume of 
material are defined by the spacing of 
the supports (width), the supporting dis- 
tance from face to gob, and the height 
and angle of caving. The height of cav- 
ing is generally estimated as a factor 
times the height of extraction, the fac- 
tor being determined by estimating the 
bulking factor of caved material. Formu- 
lae using assumed caving heights for pre- 
dicting load density vary considerably. 
Wilson (8) assumes that caved rock oc- 
cupies 1.5 times the volume of the same 
rock in situ and therefore maintains 
that the height of caving (the height of 
caving above the level of the roof) is 
twice the extraction height. Wade ( _9) , 
however, assumes that caved material will 
occupy 1.25 times the volume of rock in 
situ, yielding a caving height equal to 
four times the extraction height. 

To find the caving horizons above a 
longwall panel extraction, two instrumen- 
tation stations were installed in verti- 
cal boreholes located along the center- 
line of one longwall panel. Each station 
utilized instruments that measured both 
horizontal and vertical displacements as 
a function of longwall face advance. In 
addition, surface evaluation surveys were 
performed to differentiate between sur- 
face and subsurface displacements. 



ACKNOWLEDGMENTS 



The authors thank Roctest, Inc. ,- 
Plattsburgh, NY, for cooperation in 

— _ _ 

-"Reference to specific manufacturers 
does not imply endorsement by the Bureau 
of Mines. 



designing, manufacturing, and installing 
the borehole instrumentation. Special 
thanks is extended to Girard Theroue and 
David Prentice of Roctest, Inc., for 
their assistance at the field site. 



14 



DESCRIPTION OF TEST SITE 



The longwall panel under investigation 
is located within the Appalachian Plateau 
Province of western Pennsylvania. Struc- 
tural relief in the region does not ex- 
ceed 350 ft, and dips are generally less 
than 4°. Mining takes place in the 
Pittsburgh Coalbed, which lies strati- 
graphically within the Pennsylvania age 
coal-bearing strata of the Monongahela 
Group (fig. 1). The longwall panel is 
mined to a height of 5.8 ft; overburden 
depths vary throughout the length of the 
panel. Figure 2 illustrates the lateral 
continuity of this interval over the pan- 
el under study. The panel (panel 3) di- 
mensions are 630 ft wide and 5,570 ft 
long. This longwall panel utilizes a 
four-entry headgate and tailgate system 
with square pillars on 90-ft centers and 
entries 15 to 18 ft wide. 

Roof support along the face was main- 
tained by Dowty 460-st two-leg shield 
supports with setting pressures of 3,600 
psi. The rate of advance for the long- 
wall face was approximately 35 ft per 
three-production-shift day. 



500 
510 
520 
530 
540 
«- 550 

H— 

I 



CL 

LlI 

o 



560 
570 
580 
590 
600 
610 
620 
630 




Sewickley 
Coalbed 



Redstone 
Coalbed 



Pittsburgh 
Coalbed 



LEGEND 
I Shaly sandstone 
F^^ Shale 
■■Coal 

EJ*ii Shaly limestone 
[gig Limestone 



FIGURE 1.— Generalized stratigraphic column of the 
Monongahela Group. 



Direction of mining 

»- 

Panel 3 




1,000 



Scale, ft 



2,000 



LEGEND 



Sandstone 

Shale 

Coal 



||3jShaly limestone 
K^ Limestone 



FIGURE 2.— Fence diagram of strata overlying the study panel. 



15 



SITE PREPARATION — BOREHOLE DRILLING 



Factors considered during site selec- 
tion included topography and terrain, 
surface rights (private land ownership), 
and environmental restrictions. From a 
technical standpoint, however, the most 
important consideration was to choose a 
test site that would yield representative 
results over the length of the panel. 
Therefore, in addition to the nontech- 
nical considerations, borehole locations 
were chosen toward the center and along 
the length of the centerline of the panel 
to minimize the effects of the panel 
boundaries on the caving process. The 
two monitoring stations were located 
2,600 and 3,150 ft from the start of the 
panel (fig. 3). 

Two 6-in-diam boreholes were drilled 
through the coalbed. With the exception 
of a 15-ft standpipe at the surface, the 
holes were not cased. This allowed the 
borehole anchors to be set directly in 
distinct stratigraphic members. A combi- 
nation of rotary and core drilling was 
used to drill the first borehole (BH1). 
One hundred feet of core was extracted 
from BH1 for descriptive geologic logging 
and laboratory testing. This interval 
included the entire interburden between 
the Sewickley and Pittsburgh Coalbeds. 
The surface at BH1 had a mean sea level 
elevation of 1,004.68 ft, and the bore- 
hole depth reached 630 ft. This final 
depth was approximately 3 ft below the 
base of the Pittsburgh Coalbed. Geophys- 
ical logging was performed on each hole 
to determine lithology and the location 
of water-bearing strata. These logs were 
also used to calculate a rock strength 
index. 

The quality of the borehole is of ut- 
most importance. It is imperative to 
have a straight hole that is free of ob- 
structions so that problems during in- 
strument installation can be kept to a 



minimum. However, in this study, bore- 
hole conditions were not ideal. This 
prevented placement of borehole anchors 
at desired depths (5 ft above the coal- 
bed), so caving of the immediate roof 
behind the longwall supports could not be 
detected. Figure 4 shows a borehole di- 
rectional survey of BH1. Although the 
horizontal difference between the top and 
bottom of the borehole was only 3.5 ft, 
sharp deviations of the borehole, due to 
the spiraled drill path, prevented place- 
ment of a borehole anchor at the desired 
depth of 615 ft, which is 5 ft above the 
coalbed. 




o 

81* 



Anchor _jih 
locations! I 



h 



Anchor 
locations 



1 Pittsburgh Coalbed 



SECTION A- A' 



aooooaaooaaomaaaoaooooaoooaoDOQoao 

ODDOODDOOOaaDOQOOODQpOOOOoaoODoaDD 



BH1 



Direction of mining 



BH2 
o 



A' 



Panel 3 



aooaooaoaoCDaoaoooooaaoaoDaaaooaa^ 

oooooaaaoaooooooooooaaaoaaoaaoaaaa 

oocjaoQDaooooooooooDoaaaaooQoaoDoaa 




Scale, ft 



PLAN VIEW 



FIGURE 3.— Study panel and cross section of borehole an- 
chor locations. 



16 



ROCK MASS CHARACTERIZATION 



An assessment of geologic and mechani- 
cal rock properties of the entire unit of 
strata overlying the longwall panel at 
the mine site was conducted to relate 
these characteristics to caving height 
and lateral overburden movement. To ac- 
complish this, geophysical logging was 
carried out over the entire length of 
each of the two boreholes. 

Figure 5 is a summary of the character- 
istics obtained for the Pittsburgh and 
Sewickley interburden at the longwall 
site. The characteristics include geo- 
logic description, strength index from 
well logs (dynamic elastic modulus of 
deformation) , uniaxial compressive 
strength, indirect tensile strength, and 
rock quality designation (RQD) (11). As 
the figure shows, the immediate roof rock 
(approximately 28 ft) above the Pitts- 
burgh Coalbed is mainly composed of weak 
shales with a low RQD. Since the RQD 
values shown were calculated over each 
10-ft core run, the three clayey shale 
units (indicated as "disintegration 
zones" in the figure) do not specifically 
stand out as weak zones. However, each 
of these units disintegrated upon removal 
from the core barrel. As will be dis- 
cussed later, it is probable that caving 
height was coincident with one of these 
horizons. 




90° 
270° 



100° 
260° 



American Society for Testing and Mate- 
rials (ASTM) standards were used for 
strength evaluations of the NX core from 
BHl , including uniaxial compressive 
strength and Brazilian tensile strength 
(diametral compression) tests. All of 
the core used was ample in size to meet 
specimen standards, and a statistically 
significant number of specimens were 
tested for most of the rock types recov- 
ered. Mechanical property data could 
not be obtained for several areas in the 
borehole because either the core disinte- 
grated during recovery or the lengths of 
recovered core were inadequate for the 
preparation of test specimens. These 
areas are indicated in figure 5. 

After the hole was cored, the bottom 
100 ft was reamed to 6 in for geophysical 
logging and subsequent installation of 
the borehole instruments. Unfortunately, 
by the time geophysical logging was con- 
ducted, the tools were unable to proceed 
beyond 16 ft above the coalbed. 

The suite of geophysical tools used at 
the site included caliper, natural gamma, 
density, resistivity, spontaneous poten- 
tial, temperature, fluid conductivity, 
and sonic logs. The sonic and density 
logs were used to calculate a strength 
index for the rock above the coal seam 
where the extensometer anchors were 
positioned. 

The following equation was used to give 
a relative strength value for those sec- 
tions of the overburden that could not be 
tested in the laboratory: 



Eh = 



Pb 



x 3.36 x 10 9 



Scale, ft 

FIGURE 4.— Borehole directional survey of station 1. 



' d (6t)2 

where pb = bulk density, g/cm 3 , 

6t = interval transit time, us/ft, 

and Ed = dynamic elastic modulus of 
deformation, psi. 

This equation was developed by Schlum- 
berger Well Services (12) and relates the 
sonic and density logs to the dynamic 
elastic modulus of deformation. The 



17 



500 r 



510 



520 



530 



540 - 



550 



560 



0. 
LJ 

Q 570 



580 



590 



600 



610 



620 - 



630 L 



GE0L0GIC 
DESCRIPTION 



Stratigraphic Anchor 
column locations 



DYNAMIC ELASTIC 

MODULUS OF 

DEFORMATION, 

I0 6 psi 

I 2 



UNIAXIAL BRAZILIAN 
COMPRESSIVE TENSILE 
STRENGTH, STRENGTH, 
I0 4 psi I0 2 psi 

30 I 2 30 3 



ROCK 

QUALITY 

DESIGNATION, 

pet 

6 9 12 15 25 50 75 100 



Cross-bedded sandstone ."•'.-■■ .'-f.V" : 



Sewickley Coalbed 
Green clayey shale 
Dark-gray shale 



- Shaly limestone 



Massive limestone 

Interbedded black 
shale and limestone 

Dark-gray shale 

Massive limestone 

Interbedded green 
shale and limestone 

Massive limestone 
Green claystone 
Black shale 
Fine-grained massive 

limestone 
Clayey shale 
Green shale 
Clayey shale 

Green shale 

Clayey shale 



Black shale with 
interbedded coal 

Pittsburgh Coalbed 
Green-sandy shale 




FIGURE 5.— Geologic and rock strength characterization of borehole 1. 



value is not to be regarded as an abso- 
lute strength value but rather as an up- 
per limit of the possible strength of the 
rock. 

As compared with actual laboratory 
tests of specimens, the strength index 
showed a good correlation between low in- 
dex values and corresponding low strength 
test values. For the higher strength 
test values, the strength index often 
indicated a relatively higher strength 
than did the strength value that resulted 
from laboratory tests. Bond (12) offers 
an explanation for this relationship: 



... a competent appearing formation 
could be fractured enough to weak- 
en the rock structurally but not 
enough to create an observed effect 
on the logs. On the other hand, 
formations appearing weak on the 
strength index curve could not be 
considered stronger than the cal- 
culated index. Therefore, the 
strength index should be considered 
an indication of the upper limit of 
bed competence. 



18 



In addition to physical property test- 
ing of the immediate roof rock, in-mine 
geologic mapping was conducted in the 
gate roads of the monitored panel. All 
clastic dikes, slips, and roof falls 
were recorded as shown in figure 6. A 
high frequency of clastic dikes (often 
referred to as clay veins) is found 
throughout the study area. These dikes 
are characterized as normal fault-type 
fractures infilled with a clay matrix and 
inclusions of coal, sandstone, and shale. 
The dip of the normal fault-type frac- 
tures ranges from the vertical to 45°, 
with as much as 3.3 ft of vertical dis- 
placement along the fault plane. At this 



site, the dikes had no obvious preferred 
orientation, although their relatively 
close spacing may have facilitated cav- 
ing of the gob. Otherwise, observations 
showed that the dikes adversely affected 
ground control only in isolated areas. 

Coal cleat measurements were also taken 
at the site. The mean orientations of 
the butt and face cleat were determined 
to be N 25° E and N 65° W, respectively. 
The direction of mining was subparallel 
with the face cleat at N 60° W. Although 
no joints were found within the roof 
rock, other studies have shown that coal 
cleat orientations often mirror the ori- 
entations of joints in overlying strata 





oco 

ODOl 

OQDCD 

DQDD 

DODQ 

DDDCD 

DODC 

OOCO 

ocxr 

ODD 



.OODD 
DQQCDDDDO. 
QOQCDODDDa 



Panel 3 



March line 





SOD 

lODODQaCJO 




ClScx] 



GATE ROAD 9- RIGHT 




GATE ROAD 10-RIGHT \J 

FIGURE 6.— Roof falls and geologic anomalies of gate road entries adjacent to study panel. 



19 



( 11 , 13-14), Thus, since the face line 
of the panel was subparallel with one of 
the major orientations of the coal cleat, 
it is possible that jointing in the over- 
burden may have contributed to the caving 
characteristics of the gob. 

Overall, the rock mass characterization 
supports the favorable longwall mining 
conditions that are evident at the mine. 



The immediate roof is strong enough to 
remain stable between the tip of the 
shield line and the face, and weak enough 
to allow immediate collapse directly be- 
hind the shield line. The weak, clayey 
zones of the immediate roof appear to al- 
low for a consistent caving height, and, 
thus, consistent loading on the shields. 



INSTRUMENTATION 



Each instrumentation site was comprised 
of a surface monitoring station and a 
subsurface instrument installation as 
shown in figure 7. A premining surface 
elevation survey was performed at each 
site to establish an elevation datum, and 
successive surveys took place at inter- 
vals during panel extraction. These sur- 
veys allowed differentiation between 
surface subsidence and subsurface strata 
activity. Each hole was instrumented 
with a multiple-point borehole extensom- 
eter and an inclinometer casing, which 
allowed measurement of vertical and hori- 
zontal displacements, respectively. Ver- 
tical measurements were made by direct 
readout of extensometer scales, continu- 
ous recording units connected to the ex- 
tensometer system, and direct readout of 
a magnetic settlement probe system. Lat- 
eral displacements were calculated from 
inclinometer probe measurements. 

MONITORING OF VERTICAL 
STRATA DEFORMATION 

To measure vertical displacement of 
substrata, each borehole was equipped 
with a multiple-point borehole extensom- 
eter; this device detects vertical strata 
movement through the use of mechanical 
spring anchors. The anchors were posi- 
tioned at specific intervals within the 
strata and connected to the surface by 
stainless steel wires. Depth intervals 
for anchor placement were selected fol- 
lowing an analysis of the recovered core 
and the geophysical logs. Each anchor 
was positioned within a distinct strat- 
igraphic member. Interfaces between 
stratigraphic units were avoided on the 
assumption that caving would be most 
likely to occur along these planes. 



Pulley for lowering inclinometer 
or magnetic settlement probe 



Optional cover 



Rotary potentiometers 
and pulley assembly 



Inclinometer 

or magnetic 

settlement probe 




Spring 
anchor 



Inclinometer, 
1.9-inOD 



Centering guide 



FIGURE 7.— Complete extensometer-inclinometer system. 
(Courtesy Roctest, Inc., Pittsburgh, NY) 



20 



Each of the two 6-in boreholes accommo- 
dated eight anchors, the maximum number 
that could be used in this borehole diam- 
eter. Eight sections of 1.9-in-OD poly- 
vinyl chloride (PVC) inclinometer casing 
were prepared at the factory to accept 
the eight anchors. The remaining sec- 
tions of inclinometer casing were stan- 
dard 5-ft sections. The anchor springs 
were compressed and held closed during 
installation by nylon strings, which 
passed through the casing and were at- 
tached to opposite pairs of anchor 
springs (fig. 8). The anchors were posi- 
tioned on the casing at the predetermined 
depths shown in figure 5. The 5-ft sec- 
tions of casing were glued together and 
lowered down the hole. A grout tube was 
fastened to the first section of casing 
and lowered as the system was being as- 
sembled (fig. 9). Grouting of the bore- 
hole was necessary to seal any water- 
bearing zones, which could have caused 
water inflow into the mine when the bore- 
hole was undermined. The casing remained 
centered in the hole by means of two sets 
of PVC centering blades installed on the 
casing, 5 ft above and below each anchor 



(fig. 10). A 1/16-in-diam stainless 
steel wire surrounded by 1/4-in-diam oil- 
filled nylon tubing was attached to each 
anchor. The tubing was necessary to al- 
low free movement of the wire after the 
hole was grouted. A wire-tubing assembly 
was attached to each of the eight anchors 
and lowered with the anchor and casing 
assembly. The grout tubing and each of 
the wire-tubing assemblies were posi- 
tioned on scaffolding and lowered into 
the borehole as the 5-ft sections of cas- 
ing were added. When the entire assembly 
had been lowered into the borehole, the 
anchors were set in place by dropping a 
weighted knife down through the casing to 
cut the nylon strings. 

A reference head was placed at the top 
of each borehole. The head consisted of 
a 6-in-OD steel pipe with a welded circu- 
lar steel plate used to seat eight poten- 
tiometer-pulley assemblies (one for each 
anchor). Each anchor wire passed through 
the center of the instrument reference 
head,, over its own pulley, and was fixed 
with a 50-lb tensioning weight. Grad- 
uated scales were fastened to the outside 
circumference of the head to allow direct 





9m . 



m 



FIGURE 8.— Multiple-point borehole extensometer anchor 
positioned on Inclinometer casing. 



FIGURE 9.— Grout tube attached to lead section of in- 
clinometer casing. 



21 





FIGURE 10.— Borehole anchor and PVC centering blades. 



FIGURE 11.— Multiple-point borehole extensometer refer- 
ence head. 



readout of displacements. For remote 
readout, the potentiometer leads were 
soldered onto a terminal panel to which 
a continuous recorder was connected. The 
head also incorporated a large pulley for 
lowering the inclinometer and magnetic 
settlement probe (fig. 11). 

The magnetic settlement probe, which 
works by magnetic inductance, was used to 
verify anchor locations in the strata 
(fig. 12). Magnetic rings were incorpo- 
rated into each of the eight borehole an- 
chors, creating a magnetic field inside 
the inclinometer casing. A reed switch 
probe was connected to a graduated cable 
mounted on a cable reel. An audible buz- 
zer housed inside the cable reel was ac- 
tivated by entry of the probe into the 
localized magnetic field produced by the 
anchor. This device provided excellent 
results for verification of anchor loca- 
tions. However, after mining progressed 



beneath the borehole, high concentrations 
of methane began propagating up the bore- 
hole, preventing further use of the non- 
permissible magnetic probe and the con- 
tinuous recording unit. 

MONITORING OF HORIZONTAL 
STRATA DEFORMATION 

An inclinometer probe was used to mea- 
sure the progressive changes in the angle 
of inclination of the inclinometer cas- 
ing. These measurements provided an 
evaluation of lateral movement as mining 
approached the station. The probe was 
supported laterally in the casing by 
guide wheels and suspended vertically by 
a cable connected to a reel and readout 
unit. The guide wheels traversed oppos- 
ing longitudinal grooves spaced equally 
90° around the inside circumference of 
the casing for directional control. Two 



22 



servoacceleroraeters , mounted with sensi- 
tive axes 90° apart, simultaneously moni- 
tored inclination both parallel and per- 
pendicular to the direction of mining. 
Recording of data was accomplished by 
the use of a digital indicator equipped 
with a magnetic tape cassette recorder. 



Although the inclinometer probe output is 
recorded in terms of the angle of in- 
clination, lateral deflections can be 
calculated easily from these data. Fig- 
ure 7 illustrates the complete instrument 
arrangement. 



FIELD DATA ANALYSIS 



Base reference data were established at 
each station approximately 500 ft in ad- 
vance of mining, and readings were taken 
relative to the longwall face position. 
Surface elevation datums were also estab- 
lished for each borehole prior to mining. 
Instrument readings were taken weekly 
while the face was more than 200 ft from 
the stations, and daily when the face was 
within 200 ft of the instruments. Face 
advance was obtained from the mine daily, 
and surface elevation surveys were made 



periodically after 
the boreholes. 



mining passed beneath 







FIGURE 12.— Magnetic settlement probe. 



EXTENSOMETER DATA ANALYSIS 

To establish the progress of caving, 
borehole anchors were positioned in the 
mine roof strata as shown in figure 5. 
The deepest achor in BH1 was located 23.5 
ft above the coal seam. A distance of 10 
ft separated anchors 8 through 3, with 
anchors 2 and 1 separated by 15 and 30 
ft, respectively. 

It is important to note that the anchor 
displacements shown are with reference to 
the extensometer head located at the sur- 
face. The movements of the extensometer 
head were determined by surface elevation 
surveys. The displacements shown in fig- 
ures 13 and 15 have not been corrected 
to include the measured movements of the 
surface. 

Initial anchor movement was detected in 
all anchors when the face had approached 
to within 500 ft of station 1 (fig. 13). 
Since it cannot be associated with cav- 
ing, this movement has been attributed to 
lateral displacements of strata and to a 
rising surface elevation in advance of 
the face (fig. 14). As the face drew 
near and passed beneath the station, an- 
chor positions were recorded hourly based 
on the assumption that large movements 
would be seen immediately after the long- 
wall supports passed beneath the bore- 
hole. However, anchor movement was not 
detected at that time. Figure 13 shows 
that significant movement (caving re- 
lated) in anchors 2 to 8 began when the 
face was 35 ft past the station. Anchor 
1 (the farthest from the extraction) be- 
gan moving when the face was 75 ft past 
the station, at the same time that sur- 
face subsidence began (figs. 13-14). The 
abrupt failure of the immediate roof 



23 



Ll.1 

o 
< 
_l 

Q. 
CO 




-400 



-200 

DISTANCE, ft 



600 



FIGURE 13.— Station 1 extensometer displacements. 

associated with the advance of the long- 
wall supports was not detected by anchor 
8, which was positioned 23.5 ft above the 
extraction. This indicates that 23.5 ft 
is above the upper limit of the first 
strata separation. Descriptive geologic 
logging of the immediate strata revealed 
three very weak bands (approximately 6 in 
thick) of soft clayey shale in the imme- 
diate 25 ft of roof. These weak zones 
occurred at heights of 8, 17, and 25 ft, 
and it is assumed that immediate caving 
behind the supports occurred up to one of 
these zones. Using these zones as possi- 
ble caving horizons and relating each to 
the following equation yields three dif- 
ferent bulking factors for the extraction 
height of 5.8 ft. 

H = c + h 

H = ck 





\J.O 


1 ' 


1 ' 1 




.4 


- JK^/\ 


- 




0< 




— 


**- 


- .4 


- \ 


- 


UJ 






~ 


o 


- ,R 




k — 


•z. 








UJ 






\ 


n 








in' 


-1.2 


— 


\ — 


00 








3 








CD 


-1.6 




•k^^^ — 




-2.0 


- 


\ 




-2.4 


- 


* • - 




-? 8 


1 


1 i 1 > 



-200 200 400 600 

FACE POSITION, ft 

FIGURE 14.— Surface elevation survey data for station 1. 

where h = extracted height, ft, 

c = height of caved material from 
roof level of extracted 
height, ft, 

H = distance from floor level to 
caving horizon, ft, 



and 



k = bulking factor, unitless. 



Substitution yields 



k - - + 1. 
c 



The caving horizons of 8, 17, and 25 ft 
yield bulking factors of 1.72, 1.34, and 
1.23, respectively. The fact that anchor 
8 did not begin to detect small movements 
until the face had passed 35 ft beyond 
station 1 suggests that immediate caving 
behind the supports occurred at some 
height less than 23.5 ft. The bulking 
factor of 1.34 for the 17-ft horizon 
closely corresponds to the commonly used 



value of 1.33 for bulking of shale, which 
comprises 30 ft of the immediate roof. 
Therefore, caving is assumed to have oc- 
curred up to the 17-ft horizon. Surface 
subsidence and anchor movement occurred 
concurrently when the face was between 
75 ft and 290 ft past the borehole (figs. 
13-14). Maximum surface subsidence dur- 
ing this period was 2 ft, and maximum ex- 
tensometer movement of anchor 8 was 4.25 
in. The fact that both surface and small 
subsurface movements occurred at the same 
time and at a distance of 75 ft beyond 
station 1 suggests that the entire over- 
burden member above the assumed caved 
height of 17 ft began to sag and compact 
the gob material. 

Although the most subsidence occurred 
between 75 and 290 ft, surface movement 
did not cease until the face was 530 ft 
past the station. Anchor movement ceased 
at 290 ft. This difference is attributed 
to the closure of fractures in the strata 
above the anchors. Closure is seen as an 
apparent upward movement of anchors. 

INCLINOMETER SURVEY DATA ANALYSIS 

The inclinometer was used to detect 
lateral deflections in advance of the 
longwall face. The inclinometer probe 
measured the angle of tilt of the casing 
within the borehole in two directions, 
parallel and perpendicular to the center- 
line of the panel. The following analy- 
sis will refer to an A and B direction. 
The A direction is parallel to the direc- 
tion of mining. Positive A deflections 
are movements toward the direction of 
mining, and negative deflections are in 
the opposite direction. The B direction 
refers to deflections perpendicular to 
the direction of mining. Positive de- 
flections are movements toward the previ- 
ously mined panel, and negative deflec- 
tions show movement toward the adjacent 
unmined panel. 

The inclinometer casing in borehole 1 
was installed to a depth of 604 ft, which 
was 16 ft above the mined height of 
the Pittsburgh Coalbed. Initial readings 
were established 290 ft in advance of the 
approaching longwall face. Although the 
inclinometer casing was lowered to a 
depth of 604 ft, the initial readings 



Readings were 
approached and 
toring station, 



could only be taken to a depth of 489 ft. 
This blockage was attributed to the sharp 
deviation in direction of the borehole 
at a depth of 500 ft, as shown on the di- 
rectional survey that was performed im- 
mediately after the hole was drilled 
(fig. 4). 

taken daily as the face 
passed beneath the moni- 
but they had to be dis- 
continued after the face had passed 5 ft 
beyond the borehole because high concen- 
trations of methane were vented up the 
inclinometer casing. Inclinometer sur- 
vey data show that the vertical strata 
movements discussed in the previous sec- 
tion were accompanied by lateral deflec- 
tions of the strata. Figure 15 shows the 
variation of lateral deflection of bore- 
hole 1 relative to face advance. Figure 
15A shows movement in the A direction 
(paralle to the direction of face ad- 
vance), and figure 15B shows movement in 
the B direction (perpendicular to the di- 
rection of face advance). Figure 15 re- 
veals that shear zones begin forming at 
various depths in the borehole 263 ft in 
advance of the face. Large deviations 
from the initial reading began to occur 



I 
A, Parallel 

KEY 

Face positions 
263 ft 




100 



200 300 

DEPTH, ft 



400 



500 



FIGURE 15.— Lateral displacements of borehole 1 parallel 
and perpendicular to the direction of face advance. 



25 



when the face was 157 ft from the bore^- 
hole. In both the A and B directions, 
movement began to develop at a depth of 
112 ft. Geophysical logs show a soft 
fire clay at this depth. Two other areas 
of activity at this face position occur 
at depths of 325 and 370 ft, where the 
strata again are composed of fire clay. 

When the face advanced to within 35 ft 
of the borehole, greater movements were 
apparent. Three distinct areas are dis- 
cernible. The aforementioned depths of 
112, 325, and 370 ft show relatively 
large displacements (fig. 15). At 325 
ft, a 15-ft member of fire clay exists. 



A maximum deflection of 4.2 in can be 
seen at 360 and 480 ft. The shear zone 
begins to form at 460 ft in a 5-ft layer 
of carbonaceous shale that is situated 
between an upper member of sandstone and 
a lower member of limestone. When the 
face passed beneath the station and was 
5 ft beyond the borehole, a shear zone 
developed at a depth of 179 ft, prohibit- 
ing the probe from progressing past this 
point. A final reading taken after face 
advance had progressed more than 195 ft 
past the station revealed that the high- 
est progression of shear had extended to 
within 46 ft of the surface. 



RESULTS 



Significant anchor movement and surface 
subsidence was detected when the face had 
progressed 35 ft past station 1. Surface 
and subsurface movements continued to in- 
crease until the face had advanced to 290 
ft past the station. At this point posi- 
tive anchor deflection ceased. Surface 
subsidence continued slightly to a face 
position of 530 ft. Differential surface 
and subsurface displacement is attributed 
to the closing of fractures in the strata 
above the anchor positions. This ap- 
peared on the reference head as upward 
anchor movement. 

Anchor 8 (the deepest) in station 1 did 
not detect an abrupt failure of the roof 
immediately after passage of the supports 
beneath the borehole but moved a total of 
4.25 in as the face moved 310 ft past the 
station. This indicates that caving of 
the immediate roof occurred less than 
23.5 ft above the top of the coalbed. 
Thus, the actual bulking factor must be 
greater than 1.25. 

Three very weak bands of clayey shale 
are present between the top of the 



coalbed and anchor 8. It is assumed that 
immediate caving occurred up to one of 
these weak horizons. Calculation of a 
bulking factor based on caving to the 17- 
ft horizon yields an estimated bulking 
factor of 1.34. 

Inclinometer surveys show the formation 
of a number of shear zones throughout the 
length of the borehole. Shear zones were 
detected in station 1 263 ft in advance 
of the face. As the face drew nearer to 
the station, lateral displacements in the 
borehole became more apparent. A compar- 
ison of inclinometer survey data and geo- 
physical logs revealed that shear zones 
are associated with weak strata horizons. 
These weak horizons occurred in fire clay 
material at depths of 46, 112, 325, and 
370 ft. Another displacement of 480 ft 
occurred at a 5-ft layer of carbonaceous 
shale. As the face progressed beneath 
the borehole, a shear zone developed at 
a depth of 179 ft and prevented further 
inclinometer readings. The final shear 
zone detected was at a depth of 46 ft. 



CONCLUSIONS 



The primary goal of this investigation 
was to define the height of caving imme- 
diately behind the advancing longwall 
supports. Previous estimates of caving 
height, predict the height of caving to 
be two times (%) and four times (9) the 
extraction height. This investigation 
revealed that caving occurred less than 



23.5 ft above the coalbed, most likely 
at a height coincident with one of three 
clayey shale zones. These zones are lo- 
cated 8, 17, and 25 ft above the coalbed, 
and calculation of bulking factors based 
on caving to each of these horizons 
yields values of 1.72, 1.34, and 1.23, 
respectively. The fact that 1.34 closely 



26 



corresponds to the commonly used bulking 
factor for shale (1.33) suggests that 
the caving horizon occurred at 17 ft, or 
three times the extraction height. Re- 
sults of a field study by Matthews (_5) 
revealed similar results. 

Inclinometer data revealed another 
characteristic of longwall strata behav- 
ior. Comparison of inclinometer data 
with geophysical logs showed that the 



major lateral deflections occurred in 
weak strata (i.e. , fire clay and carbona- 
ceous shale). 

Based on these observations, the behav- 
ior of strata over longwall panels ap- 
pears to be largely dependent upon lith- 
ology. Future studies should allow the 
caving behavior of various lithologies to 
be characterized. 



REFERENCES 



1. Virginia Polytechnic Institute & 
State University. Design Optimization in 
Underground Coal Systems, Volume 10: Un- 
derground Longwall Ground Control Simu- 
lator. U.S. DOE AC01-76ET10722, Final 
Rep. , 1981, 235 pp. 

2. Kidybinski, A. Classification of 
Rocks for Longwall Caveability. Paper in 
State-of-the-Art of Ground Control in 
Longwall Mining and Mining Subsidence, 
ed. by Y. P. Chugh and M. Karmis. Soc. 
Min. Eng. AIME, 1982, pp. 31-37. 

3. Ghose, A. K. Assessment of Caving 
Characteristics and Support Resistance 
for Longwall Mining Under Massive Coal- 
measure Roof Rocks. Paper in Stability 
in Underground Mining II, ed. by A. B. 
Szwilski and C. 0. Brawner. Soc. Min. 
Eng. AIME, 1984, pp. 460-471. 

4. Bieniawski, Z. T. Rock Mechanics 
Design in Mining and Tunneling. A. A. 
Balkema, 1984, 272 pp. 

5. Matthews, J. Specifying and Ac- 
quiring Longwall Shield Supports. Paper 
in Rock Mechanics: Key to Energy Produc- 
tion, ed. by H. L. Hartraan (Proc. 27th 
U.S. Symp. Rock. Mech.). Soc. Min. Eng. 
AIME, 1986, pp. 360-366. 

6. Barla, G. B. , and S. Boshkov. In- 
vestigations of Differential Strata Move- 
ments and Water Table Fluctuations During 
Longwall Operations at the Somerset Mine 
No. 60. U.S. DOE contract ET-76-C-01- 
9041, Oct. 1978, 51 pp.; NTIS FE-9041-1. 



7. Schaller, S., and B. K. Hebble- 
white. Rock Mechanics Design Criteria 
for Longwall Mining at Angus Place Colli- 
ery. Australian Coal Industry Research 
Laboratories Ltd. , May 1981, 84 pp. 

8. Wilson, A. H. Support Load Re- 
quirements on Longwall Faces. Min. Eng. 
(London), v. 134, No. 173, 1975, pp. 479- 
491. 

9. Wade, L. V. Longwall Support Load 
Prediction From Geologic Information. 
Trans. Soc. Min. Eng. AIME, v. 262, 1977, 
pp. 209-213. 

10. Barry, A. J., 0. B. Nair, and 
J. S. Miller. Specifications for Se- 
lected Hydraulic-Powered Roof Supports. 
BuMines IC 8424, 1969, 15 pp. 

11. Deere, D. V. Technical Descrip- 
tion of Rock Cores for Engineering Pur- 
poses. Felsraech. und Ingenieurgeol. , v. 
1, 1963, pp. 18-22. 

12. Bond, L. 0., R. P. Alger, and 
A. W. Schmidt. Well Log Applications in 
Coal Mining and Rock Mechanics. Trans. 
Soc. Min. Eng. AIME, v. 250, Dec. 1971, 
pp. 354-362. 

13. Parker, J. M. Regional Systematic 
Jointing in Slightly Deformed Sedimentary 
Rocks. Geol. Soc. America Bull. , v. 53, 
1942, pp. 381-408. 

14. Ver Steeg, K. Jointing in the 
Coalbeds of Ohio. Econ. Geol. , v. 37, 
1942, pp. 503-509. 



27 



MULTIPLE-SEAM MINING PROBLEMS IN THE EASTERN UNITED STATES 
By Gregory J. Chekan, 1 Rudy J. Matetic, 1 and James A. Galek^ 

ABSTRACT 



The Bureau of Mines, in an effort to 
improve planning and development in coal 
mining, is currently investigating strata 
interactions associated with mining of 
multiple coalbeds. Strata interactions 
between adjacent coalbeds occur frequent- 
ly in the Appalachian coalfields and can 
have unfavorable effects on both product 
cost and worker safety. Two common in- 
teractions that occur between adjacent 
coalbeds are subsidence and pillar load 
transfer. At two mine sites where such 
ground interactions were present, the Bu- 
reau conducted geologic studies and gath- 
ered various geotechnical Information on 
pillar and entry stability using rock 
mechanics instrumentation. At the mine 
affected by pillar load transfer, the 
results of the study show: (1) overbur- 
den depth changed dramatically in the 
study area and reached a maximum at the 
study site, (2) innerburden thickness was 
less than a pillar width (40 to 45 ft), 



(3) heaving was experienced in both mines 
where sandstone and/or shale floor units 
were observed, (4) overlays of the mine 
layout show pillars were not totally 
superpositioned, (5) a maximum of 5 in of 
roof-to-floor convergence was measured, 
and (6) monitoring of pillar pressures 
showed that only the pillar core was 
loading, an indication of a stiff pillar 
approaching failure. 

At the mine affected by subsidence, 
measurements show that the undermining 
had little effect on upper mine pillar 
stability, but had a more severe effect 
on the development and maintenance of 
entries. Roof-to-floor measurements re- 
corded over four times more convergence 
in entries developed over gob than in en- 
tries developed over support pillars in 
the lower mine. This has led to major 
roof falls in the active panels, result- 
ing In production delays and supplemental 
roof support costs. 



INTRODUCTION 



Simultaneous mining of adjacent coal- 
beds or mining over or under a previously 
mined-out coalbed occurs frequently in 
the Appalachian Region of the Eastern 
United States. In West Virginia, Penn- 
sylvania, and Ohio, it is estimated that 
57 billion st of coal exists in a mul- 
tiple-seam configuration. West Virginia 
alone has over 50 minable seams (J_). In 
the past, mining sequence was based pri- 
marily on availability and economics with 
little regard to the effects mining would 
have on other coalbeds both above and 
below the one being mined. This has 
strong implications for resource con- 
servation, especially if these practices 



1 • • 
Mining engineer. 

Engineering technician. 

Pittsburgh Research Center, 

Mines, Pittsburgh, PA. 



Bureau of 



continue. Problems that result from 
strata interactions could render these 
resources unminable unless methodologies 
and techniques are developed that allow 
for economical extraction. 

To verify these interaction mechanisms 
in the field and their influence on mine 
ground stability, the Bureau conducted 
geological and geotechnical investiga- 
tions. The purpose of these studies is 
to develop a better understanding of sub- 
sidence and pillar load transfer and its 
effects on current workings. Eventually, 
this knowledge will lead to improvements 
in mine planning and development. 



•^Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this paper. 



28 



PILLAR LOAD TRANSFER 



Pillar load transfer, as the name im- 
plies, involves the transfer of load pil- 
lars in overlying workings to pillars in 
underlying workings. This interaction 
occurs particularly when coalbeds are in 
close proximity, less than 110 ft apart 
(2— j4) , and either isolated, remant pil- 
lars (barriers) or many strong, competent 
pillars are present in the upper work- 
ings. This condition may serve to con- 
centrate stresses in the innerburden, 
causing ground instability in the lower 
workings. The mechanics of load transfer 
and distribution from overlying opera- 
tions have been analyzed extensively by 
other researchers through the use of 
mathematical and photoelastic models (3_- 

Two theories have been developed to ex- 
plain ground disturbances due to load 
transfer from adjacent workings: pres- 
sure bulb theory and arching theory. 
Pressure bulb theory (3_-4) assumes that 
the pillar is the major structural ele- 
ment in the transfer of load. It is use- 
ful in analyzing pillar load transfer 
when a "passive" interaction occurs. 
This condition is met when pillars are 
columnized and lower seam pillars are 
sufficiently large to prevent them from 
yielding. Arching theory (3-_4) assumes 
that the mine opening is the major struc- 
tural element in the transfer of load. 
Load transfer is the result of the pres- 
sure arch that forms around the mine 
opening upon excavation. This theory is 
useful in analyzing pillar load transfer 
when a "reactive" interaction occurs. 
This condition is met when lower seam 
pillars yield and redistribute their load 
to larger barriers or abutments. 

CASE STUDY 

Mine Location and Geology 

The study mines are located in Raleigh 
County, WV, as shown in figure 1. The 
company is operating in two superimposed 
coalbeds. The upper mine is located in 
the Peerless Coalbed, which is approxi- 
mately 72 in thick. The lower mine is 
located in the No. 2 Gas Coalbed, which 
is approximately 48 in thick. The mines 



are separated by approximately 40 to 45 
ft of innerburden. 

The overburden consists predominantly 
of sandstone with interbedded shale units 
of varying thickness. The innerburden is 
a predominant sandstone with some inter- 
bedded shale units. 

Depth 

Depth in relation to all room-and- 
pillar mining operations is critical as 
overburden increases (3_-_4). The overbur- 
den above the lower mine at the study 
site is approximately 1,000 ft, the in- 
nerburden is approximately 40 to 45 ft 
thick, and approximately 960 ft of cover 
is located above the study site in the 
upper mine. Overburden depth changes 
dramatically in the area and reaches a 
topographic high over the study site. 

I nnerburden Thickness and Physical 
Characteristics 

Interval thickness between two coalbeds 
is also a critical variable. Figure 2 is 
an innerburden isopach map constructed 
from available corehole information. As 
shown on the figure, approximately 43 ft 
separates the upper and lower mines at 
the study site. Core logs show that 
sandstone comprises 77 pet of the 
innerburden. 




KEY MAP Scde ' mileS 

FIGURE 1.— Location of study mines. 



29 




LEGEND 
X Instrumented pillars Contour interval = 5ft 

O Drill holes 

FIGURE 2.— Innerburden isopach map. 





l_ 



800 



Scale, ft 



According to an equation developed by 
Haycocks and Karmis (3j-_4) , approximately 
78 ft would be the innerburden spacing 
above which no interaction damage may re- 
sult from room-and-pillar mining. The 
actual innerburden with respect to the 
study site is approximately 40 ft, con- 
siderably less than the calculated mini- 
mum value. However, it should be noted 
that the equation is independent of pil- 
lar and entry design and is derived from 
a rather limited data set and therefore 
may not represent all stable and/or un- 
stable mining conditions. 

Results of In-Mine Mapping 

A geologic investigation was performed 
within both the upper and lower mines. 
The areas of floor heave in both mines 



are underlain by a rooted, fine-grained, 
well-cemented, hard sandstone and/or 
shale, and the roof consists of either a 
hard, banded siltstone or sandstone. 
Roof structure was generally uniform ex- 
cept for some slight undulations in the 
shale-sandstone contact. The immediate 
roof rock in the heave zones of the lower 
mine appears to be sandstone. In areas 
where a shale roof was observed, no floor 
heave was detected. The heave zones in 
the upper mine could not be correlated to 
roof rock lithology since sandstone is 
present throughout. Figures 3 and 4 rep- 
resent the results of in-mine mapping. 
Figure 3 shows the location of floor 
heave around the study site in the upper 
mine. Note that the mine roof is chiefly 
sandstone near the study area and the 
floor is comprised of hard, fine-grained, 



30 



Face 



Preferred cleat 
Trends 




a 



LEGEND 

Sandstone in immediate roof 

Siltshaie roof^ hard, banded with 
ironstone (siderite) streaks and 
nodules 

Instrumented pillar 



wvw Floor heave or buckle 





1_ 



200 



Scale, ft 



FIGURE 3.— Results of in-mine mapping of upper mine. 




Buff 
sandstone 

Gray 
sandstone 

LEGEND 

^] Sandstone in immediate roof 

] Siltshaie roof-, hard, banded with 
ironstone (siderite) streaks and 
nodules 

J Instrumented pillar 

Floor heave or buckle 

Mine track 



Silt- 
shale 



200 

I 



Scale, ft 



FIGURE 4.— Results of in-mine mapping of lower mine. 



31 



rooted sandstone. Figure 5 was con- 
structed from core logs and displays 
floor lithology in the area of the study 
site. Note the lateral changes in the 
immediate floor lithology. The immediate 
floor is comprised of both a high-modulus 
material, a sandstone, and a low-modulus 
material, a shale. 

In the heaving experienced near the 
study site, there was a sandstone floor. 
Inby and outby the study site, major 
floor heaving was experienced where a 
shale floor was present. Floor heaving 
experienced in these areas was observed 
to be humplike structures characteristic 
of a low-modulus material. A buckling 
type of failure was also observed where a 
sandstone floor was present. It may be 
assumed that major concentrations of load 
can be transferred from a shale (low- 
modulus material) to a sandstone (high- 
modulus material), which may result in 
movement within the sandstone floor. 
Figure 4 represents the results of the 



in-mine mapping performed within the 
lower mine. The immediate floor located 
within the heave zones is also comprised 
of sandstone. Heave zones located in 
shale floor were found and also demon- 
strated humplike structures. The trans- 
fer of load from a shale to a sandstone 
unit could also be related to the ground 
conditions located within the lower mine. 

Mining Engineering Design Parameters 

Seam Sequencing 

The upper mine area was driven in June 
1980, and the same section located in the 
lower mine was driven during December 
1982. Major floor heaving and excessive 
pillar loading were observed in October 
1984 within the lower mine. Approximate- 
ly 3 to 4 months later the upper mine 
experienced excessive entry convergence 
and pillar loading. Affected areas are 
shown in figures 3 and 4. 



13 







20 



a. 
g o 



10- 



20 



Shale 



• Sandstone 



\ 



15 



Upper mine floor- 



Shale 



\ 



Coal 



Shale 



Shale 



Coal 



Sandstone \ 



/■\ Sandstone 
Lower mine floor 



588 



Shale 



'■'i Sandstone 



v Sandstone 



^L. 



s 



Shale 



Shale 



Sandstone 'f 



LEGEND 
15 Drill hole 
588 Depth, ft 



629 



Shale 



Sandstone 



Shale 



FIGURE 5.— Floor lithology In study site area. 



32 



Superpositioning of Pillars 

The use of coluranlzed pillars is stan- 
dard practice in multiple-seara mine 
design. Columnization of the pillars 
lessens the effect of interaction that 
may be transferred from overlying work- 
ings. Figure 6 represents the superposi- 
tioning of pillars at the study site. 
Although very difficult to achieve, this 
practice requires alignment of pillars of 
similar size for both seams. Both the 
upper and lower mines were driven with 
pillars on 70-ft centers. Note from fig- 
ure 6 that pillars and entries are near- 
ly, but not totally, superpositioned. 



stations. The pillars selected for the 
study (fig. 6) are nearly superpositioned 
and have equivalent dimensions. 

Four BPF's and 12 convergence stations 
were installed in the upper mine (fig. 
7), and 5 BPF's and 12 convergence sta- 
tions were installed in the lower mine 
(fig. 8). 

The estimated setting pressures for the 
BPF's were calculated using the tributary 
area method. 4 At the time of installa- 
tion, depth and extraction ratio were 
estimated at 720 ft and 0.35 respective- 
ly. This gives setting pressures of 
1,200 psig for the upper mine and 1,300 
psig for the lower mine. It was later 



Instrumentation and Results 



Instrumentation 



4 
(d) 



Estimated pressure, psi = 1.1 psi/ft 
1 
(1-R) 



Instrumentation installed in both mines 
included borehole platened flatjacks 
(BPF) (_5) and removable convergence 



where d - depth, ft, 

and R = extraction ratio. 




LODDQD 



D 
G 

a 
a 



LEGEND 

Instrumented 

pillar 

Upper mine 
Lower mine 



1 00 



Scale, ft 
FIGURE 6.— Superpositioning of instrumented pillar. 




LEGEND 
o Borehole platened flatjack 
• Convergence station 



40 



Scale, ft 
FIGURE 7.— Instrument location in upper mine. 



33 




LEGEND 
o Borehole platened flatjack 
• Convergence station 



40 

_l 



Scale, ft 



FIGURE 8.— Instrument location in lower mine. 

determined that depth and extraction ra- 
tio were underestimated and were actually 
960 ft and 0.50. As a result, setting 
pressures for the upper and lower mines 
should have been 2,100 psig and 2,200 
psig respectively. Although the original 
setting pressures were low, these pres- 
sures do not directly affect the recorded 
results. Any increase in pillar pressure 




MMJMt Ml j 1 ! 1 



20 40 60 80 100 120 140 160 180 

TIME, days 
FIGURE 9.— Pressure increase versus time in upper mine. 

above 2,100 psig and 2,200 psig would be 
a result of relative increases in pillar 
pressure. 

Results 

The monitoring of the instrumentation 
continued for a total of 177 days. The 
instruments were monitored at least once 
a week. Tables 1 and 2 provide results 
at 49 days (approximately 25 pet of the 
study period), 92 days (approximately 50 
pet of the study period), and 177 days. 
All increases in pillar pressure for the 
upper mine were from BPF's installed 
within the core of the instrumented pil- 
lars. The BPF's located at 10-ft depths 
showed no increase in pressure. Figure 9 
shows pressure increases versus time for 



TABLE 1. - Flatjack (BPF) pressure during 177-day 
monitoring period, pounds per square inch 



BPF 


Initial (instal- 


Day 49 (25 pet 


Day 92 


Day 177 




lation date) 


of total period) 


(50 pet) 


(final) 


UPPER MINE 




1,100 


8,100 


8,100 


8,100 




1,225 


950 


900 


950 


3 


1,200 


3,175 


4,050 


5,100 


4 


1,275 


1,200 


1,250 


1,250 


LOWER MINE 


5 


1,300 


1,090 


1,050 


1,050 


6 


1,000 


880 


850 


850 




1,300 


1,050 


1,000 


1,050 


8 


1,300 


900 


900 


900 


10 


1,200 


1,000 


1,000 


950 



34 



TABLE 2. - Results of convergence monitoring in upper mine, inches 



Station 


Day 49 


Day 92 


Day 17 7 
(final) 


Station 


Day 49 


Day 92 


Day 177 
(final) 


1 


1.40 

.84 

ND 

.60 

1.24 

1.32 


1.95 
1.25 
ND 
1.29 
1.88 
2.03 


3.54 

2.13 

ND 

U.so 

2.71 
3.37 


7 


1.15 
.81 

2.23 
.94 
.80 

1.40 


1.85 
1.26 
3.26 
1.31 
1.13 
1.45 


3.00 


2 


8 


2.08 


3 


9 


5.00 


4 


10 


2. 12 


5 


11 


1.92 


6 


12 


2.24 



ND No data because station was destroyed. 

Monitoring discontinued on day 134 owing to bad roof conditions. 



all BPF's installed in the upper mine. 
No increases in pillar pressure were re- 
corded from BPF's installed in the lower 
mine. This was due to height restric- 
tions which limited instrument installa- 
tion to more stable mine areas. 

Maximum convergence of 5 in occurred at 
station 9, which is located in the track 
entry within the upper mine. Roof-to- 
floor convergence within the upper mine 
increased very rapidly, averaging 0.5 in 



of closure every month. Figures 10 and 
11 show total convergence within the up- 
per mine after 49 days (25 pet of moni- 
toring period) and 177 days (100 pet of 
monitoring period). Note the trend of 
movement with respect to figure 11. 
Movement occurred outby and in a south- 
west direction in relation to the instru- 
ment array. Figure 12 displays total 
convergence versus time for the four con- 
vergence stations located along this 





LEGEND 40 

-— -1.5 — Convergence contour line; Sca | e fr 
contour interval 0.5 in 

• Convergence station 

FIGURE 10.— Upper mine convergence contours after 49 
days of monitoring. 




LEGEND 



40 



—-I-5 — Convergence contour line; Scale, ft 
contour interval 0.5 in 
• Convergence station 

FIGURE 11.— Upper mine convergence contours after 177 
days of monitoring. 



35 



trend of major roof-to-floor movement. 
Roof-to-floor convergence monitored in 
the lower mine was limited. The most 
movement that occurred was 0.13 in at 
station 11. Figure 13 represents total 
convergence (117 days) monitored within 
the lower mine. 



UJ 

o 

z 

UJ 

o 
cr 
ui 
> 

z 
o 
o 



3 - 



2 - 



I - 



KEY ^^ 


Station: ^^"^ 

• 10 -^ A 


j^**^ 



20 40 60 80 100 120 140 160 180 

TIME, days 

FIGURE 12.— Total convergence versus time in upper mine. 




LEGEND 

-O.i — Convergence contour line; 
contour interval 0.5 in 

• Convergence station 



40 

I 



Scale, ft 



FIGURE 13.— Lower mine convergence contours after 177 
days of monitoring. 



CONCLUSIONS ON PILLAR LOAD TRANSFER 

Based on the information received and 
collected throughout the study, the fol- 
lowing conclusions can be made: 

1. Overburden depth above the study 
site was approximately 1,000 ft. Prior 
research and other case studies (_2-4) 
have shown that excessive overburden 
depths can lead to unstable ground 
conditions. 

2. Innerburden thickness in the study 
area was approximately 40 to 45 ft, less 
than one pillar width. Prior research 
has shown (3-4) that workings in close 
proximity, less than two pillar widths, 
may create ground control problems in, 
above, and below workings. 

3. Prior research has shown (3-_4) that 
innerburden material comprised mostly of 
sandstone dampens the effects of pillar 
load transfer. Sandstone at the study 
area comprises 77 pet of the innerburden. 
According to Haycocks (3) this percentage 
requires 78 ft of innerburden for stable 
conditions. 

4. Heaving was experienced in both 
sandstone and shale floor units. The 
shale floor was a low-modulus material, 
resulting in humplike floor heaving. 
Whereas the sandstone floor was a high- 
modulus material, resulting in a buckling 
type of floor heave. 

5. Average convergence in the upper 
mine entries was 2.50 in, compared with 
0.04 in. in the lower mine entries. This 
difference is due to height restrictions 
in the lower mine, which limited instru- 
ment installation to a more stable mine 
area. 

6. To minimize interaction effects, 
optimum seam sequencing would be to mine 
the upper seam first to total extrac- 
tion and then continue downward. In this 
case, the upper seam pillars were devel- 
oped first, and the lower seam pillars 
were developed approximately 2 yr later. 
Pillar columnization was practiced, but 
mine overlays show that pillars and en- 
tries were not totally superpositioned. 

It is difficult to determine whether 
columnized pillars cause a "passive" in- 
teraction through pressure bulb inter- 
ference or if a "reactive" interaction 



36 



occurred owing to arching effects. BPF 
pressure readings in the upper mine show 
a core loading that is characteristic of 
a stiff pillar approaching failure. Sim- 
ilar loading characteristics were not ob- 
served in lower seam pillars, mainly be- 
cause height restrictions limited BPF 
installation to more stable areas. But 
assuming lower seam pillars also exhibit 
stiff pillar characteristics, the arch- 
ing concept could be applicable in this 
situation. Photoelastic and mathemati- 
cal models of multiple openings in close 
proximity, less than two pillar widths, 
have shown that pressure arch interaction 
can create zones of excessive pressure in 
the innerburden (3-4). This is especial- 
ly the case for strong, competent pil- 
lars or stiff pillars that do not yield 



readily, allowing independent arches to 
form from pillar to pillar. When high 
abutment pressures associated with arch 
interaction exceed the in situ strength 
of the rock, ground failure results. In 
this case, a soft floor stratum is the 
weakest member in the mine structure. 
Strong competent pillars punching into 
the floor cause considerable floor heave, 
which is then followed by eventual pillar 
failure or yielding. When this occurs, 
their load is transferred to neighboring 
pillars, forming a secondary arch. This 
cycle of failure, load transfer, and arch 
formation continues until sufficient sup- 
port is encountered to stabilize the load 
transfer process, such as barrier or 
abutment pillars. 



SUBSIDENCE 



Strata interactions due to subsidence 
result when an underlying coalbed is ex- 
tracted first. Undermining subjects the 
superjacent strata to a mining-induced 
stress field (6_). The distribution of 
this stress within the strata is a func- 
tion of the subsidence process and is 
most damaging to overlying coalbeds after 
the critical to supercritical subsidence 
phase has been reached (4_, _6_-^7). Depend- 
ing upon the uniformity of lower coalbed 
extraction, there exists a relatively de- 
stressed zone toward the middle of the 
subsidence area. Most ground distur- 
bances in overlying coalbeds occur toward 



_ Outer limit 
of subsidence 



,\ § j<> 



Tension zone 
(roof cracking and failure) 

/ Roof 

.■>"">*/ ^compression zone 



Dead ground 




Unsubsided ground 



Subsidence zone 



FIGURE 14.— Strata flexure in upper coalbed due to sub- 
sidence. Adapted from Haycocks, Karmis, and Topuz (7). 



the boundaries of the subsidence trough. 
Within the trough, strata flexure creates 
zones of tensile and compressive stress, 
as shown in figure 14 (17). The extent 
of this zone is defined by the angle of 
draw, which is dependent upon the geo- 
logic and physical characteristics of the 
strata. As mining develops through this 
trough, these stresses have a severe ef- 
fect on entry stability, particularly on 
the integrity of the roof. 

Other types of failure in upper coal- 
beds that are attributed to undermining 
include interseam shearing and the ef- 
fects of arching. In interseam shearing, 
highly inclined shear or tensile-shear 
failures develop and result in displace- 
ment of the coalbed into lower excava- 
tions (4_, 7). Recent studies (4_, 7) in- 
dicate that the elastic modulus of the 
superincumbent strata is a major factor 
influencing this type of failure. Stud- 
ies have demonstrated that a high-modulus 
strata such as sandstone is more prone 
to shear failure. Arching is actually a 
subcritical phase of subsidence, and its 
effects on upper coalbeds are dependent 
upon the opening width-to-depth ratio and 
the height of the resulting pressure 
arch. Arching effects can produce a zone 
of high compressive stress that may cause 
pillar and roof control problems. 



37 



CASE STUDY 
Mine Location and Geology 

The study mine is located in Greene 
County, PA, as shown in figure 15 and is 
operating in areas of the Sewickley Coal- 
bed that were subsided by undermining of 
the Pittsburgh Coalbed. The overburden 
above the Sewickley Coalbed ranges from 
425 to 580 ft and consists predominantly 
of interbedded shale with a competent 
sandstone unit that varies in thickness. 
The innerburden ranges from 90 to 100 ft 
and consists of interbedded shale and 
limestone. 

The average height of the Sewickley 
Coalbed is 5 ft. The immediate roof is 
composed of a highly jointed dark sandy 
shale that ranges from 10 to 15 ft thick, 
overlain by a competent limy shale. The 
immediate floor is composed of a 3-ft- 
thick, dark, limy shale underlain by a 
competent limestone unit. An underground 
geologic investigation found no geologic 
anomalies (clay veins, discontinuities, 
etc. ) in the study area. Additional 
site-specific information is given in 
table 3. 

Instrumentation and Results 

1 Left Panel 

1 Left was a short 400-ft panel that 
was developed through a subsided area of 
the coalbed created by pillar retreat ac- 
tivities in the lower mine. The boundary 

TABLE 3. - Site-specific information, 1 
Greene County, PA, mine 





Upper 
mine 


Lower 
mine 3 


Av mining height.. 
Pillar centers.... 
Av entry width. . . 
Extraction, pet: 


• in. . 

.ft. . 

, .ft. . 


60-64 

100x100 

18-20 

36 

100 


72 

100x100 

20 

36 




100 



^oom-and-pillar continuous mining at 
both levels. 

Sewickley Coalbed. 
Pittsburgh Coalbed. 




GREENE 
COUNTY 



Pennsylvania 
key map 



Pittsburgh 



Umontown 
PENNSYLVANIA 



MARYLAND 



Morgantown 



Scale, mi 

FIGURE 15.— Study mine location, Greene County, PA, 
mine. 



of this retreat raining in the Pittsburgh 
Coalbed is shown as the gob line in fig- 
ure 16. The 1 Left panel started in sub- 
sidence (over gob) and developed across 
the gob line and onto pillars located in 
the lower mine. It was developed and re- 
treated in less than 50 days. Several 
years earlier, the 1 East panel crossed 
the gob line (from over support pillars 
and onto gob in the lower mine); it ex- 
perienced two roof falls as shown in fig- 
ure 16 but encountered no displacements 
in the coalbed. These ground conditions 
indicated that the strata directly super- 
jacent to the gob line were flexed due to 
subsidence, and mine personnel were an- 
ticipating these same conditions in the 
1 Left panel. To predict the location of 
the subsidence trough within the coalbed 
is a complex problem. Although current 
theories and models are based on surface 
subsidence, their application for pre- 
dicting in-mine subsidence could prove 
useful. Based on a model recently devel- 
oped by the Bureau (8-9) , the edges of 
the subsidence trough at the Sewickley 
Coalbed level were calculated to be ap- 
proximately 102 ft inby to 300 ft outby 
the gob line of the lower mine for this 
case. The edges of the predicted trough 
and pillar arrangements in 1 Left before 
retreat mining are shown in figure 17. 



38 



DDDDDD 
□DQDDD 

DDDDnC 
DDDQDD 

□DODOD 

DDDDnDc 

DDDDDDSmDr 



I East Panel 



ODD 



□DsnnDnrag! 



DDD 

nan 



DDD 
DDD 



DDI 



□□DM 
DaQDOTa 
□□□DDD 



DDI 
DDD 

DDDDBa 

aDDaai 




nan 

□CL1DD 



LEGEND 
■ Convergence station 
cOd Roof falls occurring in I East 
when mining crossed the 
gob line 
GZg Roof falls occurring in I East 
since mining was discontinued 
in September 1984 

•» Roof falls occurring in 2 East 
during development 

Gob line - lower mine 





















t 











2 East Panel 



200 




Scale, ft 



FIGURE 16.— Location of study areas 1 Left and 2 East showing gobline in lower mine (Pittsburgh Coalbed) and roof fall activity 
in upper mine (Sewickley Coalbed). 



To determine if the trough affected 
pillar loading and stability, four BPF's 
(12) were installed during development to 
measure vertical changes in pillar pres- 
sure. Setting pressure was approximately 
1,000 psig for each BPF 5 as overburden 
was 580 ft with an extraction ratio of 
0.36. 

Figure 17 shows BPF locations in se- 
lected pillars. During development the 
instruments recorded no increases in pil- 
lar pressure. The major ground problem 
experienced within the trough during de- 
velopment was the occurrence of a large 
roof fall, 10 to 12 ft high, over gob as 
shown in figure 17. This fall started in 
a crosscut at the face and eventually 
propagated into the belt entry, causing 
considerable production delays. 

BPF pressure increased, as anticipated, 
when retreat raining approached the in- 
strumented pillars. Figure 18 shows the 
pressure changes recorded for the four 
BPF's during development and retreat 



5 See footnote 4 (p. 32). 



mining. Pressure increases ranged from 
400 to 700 psig but were not considered 
significant because they did not render a 
pillar or adjacent pillar unminable dur- 
ing retreat. Observations of roof behav- 
ior at the face showed the roof strata 
to fall tight against the pillar line, 
causing no excessive loading in adjacent 
pillars from roof cantilevering. The 
1 Left panel was completely retreated, 
but ground water inflow and accumulation 
slowed retreat operations in the 1 East 
panel. Eventually, retreat mining was 
discontinued completely in the 1 East 
panel due to this condition. 

2 East Panel 

2 East was a long 2,200-ft panel devel- 
oped off the 1 South Mains. It started 
from over support pillars located in the 
lower mine and developed across the gob 
line into subsided ground as shown in 
figure 16. To better understand immedi- 
ate ground movement as entries were de- 
veloped through the subsidence trough, 



39 



I 




T* 


r 



□E 






ra 



I East Panel 
(direction of mining) 



16 

15 

.? 14 

w 

a. 



CVJ 



O |3 



LU 



o 

LU 
CC 

CO 
CO 

UJ 

or 



12 



10 







1 1 

KEY 
- BPF No. 
A | 


1 1 

x x — 


■ 2 




• 3 




A 4 


/ / x 


- x BPF lost 


/ // ~~ 


to retreat 




_ mining 


/ 1/ ~ 


/• BPF 
/installation 


x a y 


/V pressure 


i / ~ 


X^ \ Y^_ ^LX 


__^j ^r 




.--- Start of retreat 


1 


^^ mining _ 

1 1 



10 



20 30 

TIME, days 



40 



50 



FIGURE 18.— Pressure changes recorded for BPF locations 
in 1 Left before development and retreat mining. 



I oo 



200 



LEGEND 
o Borehole platened flatjack o 

eZZ) Roof fall ' i J^ ' 

— — Gob line- lower mine 
Edges of predicted sub- 
sidence trough 

FIGURE 17.— Pillar arrangements and BPF locations in 1 
Left before retreat mining. 



roof-to-floor convergence measurements 
were used in the 2 East panel. A total 
of 25 convergence stations were installed 
within the trough area, 14 in entries 
developed over support pillars and 11 in 
entries developed over the gob of the 
lower mine. As shown in figure 16, large 
roof falls 10 to 12 ft high occurred 
over gob during pillar development, but 
with greater frequency than in other gob 
line crossings. Convergence measurements 
taken over 143 days showed that the aver- 
age total convergence measured in entries 
developed over the gob was over four 
times the amount measured in entries de- 
veloped over support pillars. Figure 19 
shows cumulative convergence for sta- 
tions 1, 8, and 13 installed in entries 



2.0 




80 
TIME, days 



160 



FIGURE 19.— Cumulative convergence for stations 2, 8, 13, 
18, 22, and 25 over 143 days. 



developed over support pillars and sta- 
tions 18, 22, and 25 installed in entries 
developed over gob. This graph shows 
that the rate of convergence is nearly 



40 



uniform on both sides of the gob line but 
increases dramatically once the gob line 
is crossed. Table 4 lists cumulative 
convergence measured over 143 days for 
all stations. 

CONCLUSIONS ON SUBSIDENCE 

Throughout the subsidence area, the 
coalbed showed no vertical displacements 
that would indicate an interseam shear- 
ing, and no recurring problems where a 
result of strata flexure. Several obser- 
vations support this: 

First, as mining development crossed 
the gob line in the 2 East panel, roof- 
to-floor convergence increased, as did 

TABLE 4. - Cumulative convergence 
measured over 143 days for all 
stations 

Cumulative 
Station convergence, in 



0.11 
.15 
.16 
.13 
.19 
.12 
.15 
.24 

1.39 
.27 
.16 
.17 
.31 
.95 



.33 

.96 
1.11 
1.56 
1.50 
2.71 
1.46 
1.38 
1.38 

.73 
1.33 



Entries developed 

support pillars: 

1 


over 


2 


3 


4 


5 


6 


7 


8 


9 


10 


11..., 


12 


13 


14. 


Entries 
gob: 
15 


developed 


over 


16 


17 


18 


19 


20 


21 


22 


23 


24 





the incidence of roof falls. In figure 
20, graph 1 is the predicted subsidence 
profile (_8~9) for this case, graph 2 
shows cumulative convergence measured 
over 143 days for selected stations 2, 8, 
13, 18, 22, and 25 and their locations 
with respect to the gob line, graph 3 de- 
picts the cumulative length of roof falls 
with respect to the gob line, and graph 4 
Is a plot of the overburden above the 2 
East panel. Graph 1 shows that the sub- 
sidence trough begins 102 ft outby the 
gob line and subsidence reaches a maximum 
at 300 ft inby the gob line. Graphs 2 
and 3, which depict ground movements, 
correlate well with this predicted subsi- 
dence profile as roof-to-floor conver- 
gence and frequency of roof falls both 
increase dramatically as the gob line 
is crossed. Graph 4 shows there were no 
dramatic fluctuations in overburden above 
the panel. Second, all roof falls in the 
study area occurred in entries developed 



500 



G *1 

= H 450 

CO q_ 

° 400 



1 ■ ' 1 


• 1 


4 


. Direction ot mining ~^ 

in 2 East / 1 


"^\ 


■*" ^S ' 


y 


- 


^^ 


1 


- 




2 in 

3 > 

o -z. 
o 
o 



Q - 
UJ bJ 



1 


i ■ ' i 

KEY 

18 Station 1 

- ■■► 

1 

2 8 Hh 


18 

I 




22 

I 




1 
25 

I 


1 I 

2 



or w 
o_ no 





I 

2 

3 

4 
-600 



- 


1 \ 


i i ' 
/ 


- -► 


\ 

Gob line H \ 

1 N 


- 


' ' 


i , , i 


\_ s MAX 


i 



-300 300 

DISTANCE FROM GOB LINE, ft 



600 



FIGURE 20.— All graphs in relation to gobiine of the lower 
mine. 1, Predictive subsidence; 2, cumulative convergence; 3, 
cumulative length of roof falls; 4, overburden depth. 



41 



over gob; no roof falls occurred in en- 
tries developed over support pillars. In 
addition, the roof fall activity contin- 
ued in the 1 East panel well after re- 
treat raining operations were discontinued 
in September 1984 (fig. 16). 

Finally, most roof falls displayed a 
similar type of roof failure, usually 
along a natural roof joint. As shown in 
figure 21, these joint surfaces were 
smooth, having no cohesive properties. 
This natural jointing system was present 
throughout the study area, yet as men- 
tioned earlier, all roof falls occurred 
in entries developed over gob. Presuming 
that the roof is in flexure, these joint 
surfaces would provide natural planes 
for tension failure. A major set of 
roof joints was oriented approximately 
N 55° W, subparallel to the direction of 
mining. This could further explain the 
high frequency of roof falls in intersec- 
tions and entries parallel with the face. 




The increased incidence of roof falls 
within the subsidence trough required the 
installation of additional roof support. 
Roof stability in entries developed over 
support pillars was readily maintained 
with 6-ft conventional bolts on 4-ft cen- 
ters. Cribs and posts were installed as 
dictated by the general roof support 
plan. Entries developed over the gob re- 
quired more comprehensive roof support 
consisting of 5-in by 7-in by 16-ft tim- 
bers bolted on 2-ft centers. In addi- 
tion, many intersections along the track 
and belt entries were supported with 6-in 
steel I-beams set on posts or cribbing. 
These support requirements, combined with 
downtime to clean and resupport fall 
areas, lowered production considerably. 
Figure 22 shows the average tonnage per 
shift as the 2 East panel mined through 
this subsidence trough versus tonnage for 
a similar panel, 2 North, located in the 
same mine but not affected by subsidence. 
Note that the high production values for 
the 2 East panel are less than the low 
values for the 2 North panel. As raining 
progressed farther into the subsided 
zone, roof conditions improved slightly 
and production increased, but this sup- 
plemental support was still required. 

Based on information collected at this 
particular study site, the following con- 
clusions can be made: 



400 



300 



X 
CO 



W 200 



O 

O 

Q 100 

o 
<r 
a. 



KEY 



I I 2 East panel 
| 2 North panel 




JAN. 



FEB. 



MAR. 



FIGURE 21.— Failure along a roof joint. 



FIGURE 22.— Production figures, January-March 1985 - 2 
East versus 2 North. 



42 



1. Subsidence in the Sewickley Coal- 
bed, caused by undermining of the Pitts- 
burgh Coalbed, resulted in lower produc- 
tion and increased roof support costs 
during development and retreat mining. 

2. Entry convergence and roof fall 
activity increased dramatically as devel- 
opment crossed the gob line and proceeded 
into subsided ground. Falls occur pre- 
dominantly in highly jointed zones in the 
roof strata. The direction of mining in 
relation to the direction of roof joints 
may contribute to roof fall activity. 



3. Strata flexure has little effect on 
pillar stability when percent extraction 
is kept low, leaving sufficient coal in- 
place. In this case, pillars were 80- by 
80-ft with 36-pct extraction. 

4. The predicted location of the sub- 
sidence trough correlated well with re- 
corded ground movements and observed 
ground behavior. The application of this 
model for predicting in-mine subsidence 
will be studied further at this and other 
sites. 



DISCUSSION 



Interactions between adjacent workings 
that are a result of pillar load transfer 
are difficult to predict, for the complex 
nature of the mechanism is not yet fully 
understood. Once the mechanism respon- 
sible for this interaction is better 
defined through further laboratory and 
field research, problem areas may be an- 
ticipated and possibly eliminated. Ini- 
tial research indicates that the pressure 
arch theory is a feasible concept in un- 
derstanding the interaction mechanism but 
requires further study. The literature 
reports in many cases that pillar super- 
positioning and a mining sequence of up- 
per coalbed first can increase the sta- 
bility of lower workings, but this is not 
always entirely effective. Barriers and 
abutment pillars are also beneficial in 
limiting interaction effects by providing 
immediate support to load transfer. 
Problems associated with subsidence of an 
overlying coalbed can be reasonably pre- 
dicted through careful premine planning 
and analysis. This should include accu- 
rate maps of lower mine workings, a geo- 
logic evaluation of innerburden and over- 
burden, and the mapping of roof joints 
and geologic features. This informa- 
tion together with other site specifics 
(age of workings, uniformity of pillar 



extraction, mining height, etc.) should 
be used collectively to anticipate areas 
within the coalbed where problems may 
result or be magnified by undermining. 
Continued research into the development 
and application of a subsidence model 
will improve the accuracy of prediction. 

Minimizing problems may require chang- 
ing the normal mining or roof control 
plan. Some possible changes in the min- 
ing plan may include reducing percent ex- 
traction, decreasing roof span width, 
staggered pillar arrangements, and chang- 
ing the heading of entries. The roof 
control plan may also require modifica- 
tion to contend with changing roof condi- 
tions. This may require different roof 
bolt types and lengths as well as in- 
creased supplemental supports. 

Owing to economics, ownership, and 
availability, a mine operator may have 
little control over coalbed mining se- 
quence. In those cases where underlying 
coalbeds are extracted first, subsidence 
in upper beds may result in lower produc- 
tion and increased roof support costs. 
Cooperative planning within the mining 
community can avoid this and lead to 
substantial improvements in resource 
conservation. 



REFERENCES 



43 



1. Singh, M. A., and M. F. Dunn. 
Investigation of Problems and Benefits of 
Underground Multiple Seam Mining (U.S. 
DOE contract DEAC01-79ET14242, Engi- 
neers International, Inc.). Aug. 1981, 
292 pp. 

2. Stemple, D. T. A Study of Problems 
Encountered in Multiple-Seam Coal Mining 
in the Eastern U.S. Bull. VA Polytech. 
Inst., v. 49, No. 5, Mar. 1956, 65 pp. 

3. Haycocks, C. , B. Ehgartner, M. Kar- 
mis, and E. Topuz. Pillar Load Trans- 
fer Mechanics in Multi-Seam Mining. Soc. 
Min. Eng. AIME preprint 82-69, 1982, 
7 pp. 

4. Haycocks, C. , M. Karmis, E. Barko, 
J. Carmen, B. Ehgartner, S. Hudock, and 
S. Webster. Ground Control Mechanisms 
in Multi-Seam Mining (grant G1115511, 
VA Polytech. Inst.). BuMines OFR 7-84, 
1983, 328 pp. 

5. Bauer, E. R. , G. J. Chekan, and 
J. L. Hill III. A Borehole Instrument 
for Measuring Mining-Induced Pressure 
Changes in Underground Coal Mines. Paper 
in Research and Engineering Applica- 
tions in Rock Masses, ed. by E. Ashworth 
(Proc. 26th U.S. Symp. on Rock Mech. , SD 
Sch. Mines and Technol. , Rapid City, SD, 



June 26-28, 1985). A. A. Balkema, 1985, 
pp. 1,075-1,084. 

6. King, H. J., B. N. Whittaker, and 
A. S. Batchelor. The Effects of Interac- 
tions in Mine Layouts. Paper in Proceed- 
ings of the Fifth International Strata 
Control Conference. Nat. Coal Board, 
Paper 17, 1972, 11 pp. 

7. Haycocks, C. , M. Karmis, and E. To- 
puz. Optimizing Productive Potential in 
Multi-Seam Underground Coal Mining. Pa- 
per in Symposium on Underground Mining 
(Proc. Coal Conf. and Expo VI, Louis- 
ville, KY, Oct. 27-29, 1981). McGraw- 
Hill, 1981, pp. 151-163. 

8. Adamek, V., and P. W. Jeran. Eval- 
uation of Surface Deformation Character- 
istics Over Longwall Panels in the North- 
ern Appalachian Coal Field. Paper in 
Proceedings of the International Sympo- 
sium on Ground Control in Longwall Coal 
Mining and Mining Subsidence - State- 
of-the-Art (Honolulu, HI). AIME, 1982, 
pp. 183-197. 

9. . Precalculations of Subsi- 
dences Over Longwall Panels in the North- 
ern Appalachian Coal Fields. Soc. Min. 
Eng. AIME preprint 85-404, 1985, 11 pp. 



44 



INTEGRATED DESIGN FOR STABILITY IN MULTIPLE-SEAM MINING 
By Chris Haycocks, 1 Wei Wu, 2 and Yingxin Zhou 2 



ABSTRACT 



Research into ground control problems 
resulting from mining in a multiple-seam 
environment has been carried out using 
statistical analysis, numerical modeling, 
and body-loaded multilayer photoelastic 
analysis in conjunction with numerous 
case studies. Findings demonstrate situ- 
ations under which the worst ground con- 
trol conditions can be expected to occur 
for a specified geologic environment, 
depth, innerburden spacing, and mining 
geometry. Special emphasis has been 
placed on the effect of upper pillar load 
transfer on the lower seam in terras of 



pillar and floor stability and roof con- 
trol. Geologic and lithologic conditions 
in the affected seam are summarized by 
roof and floor index. The effects of 
joint orientation and continuity are also 
incorporated in the analysis. Upper seam 
effects due to undermining are given in 
terras of roof damage related to total 
vertical displacement with allowances for 
geologic variations. Design criteria for 
mining on a lower seam have been incorpo- 
rated into a friendly interactive program 
for use on personal computers. 



INTRODUCTION 



The majority of Appalachian coal depos- 
its lie in a multiple-seam environment. 
Historically, the decision to extract a 
particular seam has been based on ease of 
access, ownership, and economics rather 
than ground control considerations. As 
mining progresses and seams are mined 
out, it is increasingly common to find 
working and abandoned mines in close ver- 
tical proximity to each other. Depend- 
ing on the sequence of extraction, such 
multiple-seam mining operations can be 
classified in terms of mining sequence as 



undermining, overmining, or simultaneous 
mining. Contiguous operations can pro- 
duce both positive and negative ground 
control situations in neighboring seams. 
These interaction effects are frequently 
reported, and their severity can range 
from improved opening stability to minor 
roof problems that may require no addi- 
tional support to complete loss of a 
large section of coal reserves (l_-4_). 3 
As a consequence, interaction problems 
are of growing concern to the Appalachian 
underground coal industry. 



FIELD STUDIES 



Since many Appalachian coal mines have 
experienced interaction problems (_L - .2> 
_5 ) , there are considerable case studies 
available for research and analysis. To 
date, over 130 separate examples of mul- 
tiple-seam ground control problems have 
been collected to facilitate identifica- 
tion of major controlling factors and 
the magnitude and limits of interaction 
under a variety of conditions. 

Professor of mining engineering. 
^Graduate research assistant. 
Virginia Polytechnic Institute and 
State University, Blacksburg, VA. 



Analysis of field experiences reveals 
four major classes of interactive ground 
control mechanisms: pillar load trans- 
fer, innerburden shearing, arching ef- 
fects, and upper seam subsidence (3, 6). 
Pillar load transfer mechanisms and in- 
nerburden shearing have been successfully 
used to explain interaction phenomena 
associated with undermining during 
overmining. 

3 Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this paper. 



45 



Factors that contribute to interaction 
may be classified into variables that are 
fixed by the geologic environment and 
those that depend on engineering design. 
These may be described 4 as follows (6^): 

Fixed: Depth, innerburden thickness, 
innerburden physical characteristics,' 
stress field, seam thickness, coal char- 
acteristics, completed mining operations 
either above or below the seam to be 
mined, and age of workings. 

Mining: Height, dimensions and geome- 
try, method, and spatial location of 
entries. 

Sorting of data revealed that, consid- 
ering all case studies, 46 pet reported 
relatively serious interaction problems. 
Among the mines experiencing interaction 
problems, simultaneous mining caused the 
most trouble and accounted for 48 pet 

^Factors marked with an asterisk are 
regarded as being critical. 



300 
280 - 
260 - 
240 
C220 
$~200 

UJ 

I 180 

E 160 



uj 140 

Q 

cr 

CO 

uj !00 



20 



80 
60 
40 
20 







KEY 
+ Unstable 
• Stable 






_L 



10 20 30 40 50 60 70 80 90 100 110 
INNERBURDEN SANDSTONE, pet 



FIGURE 1.— Influence of percentage of hard rock in the in- 
nerburden on stability in the lower seam. Adapted from 
Haycocks (3). 



of the problems; overmining and under- 
mining accounted for 30 and 22 pet, 
respectively. 

Initial analysis of field study data by 
plotting the data with respect to some 
related variables also demonstrated some 
simple but useful relationships among 
variables and some interesting trends 
(figs. 1-3). 



320 



_ 280- 
<+- 

$240 

UJ 

g 200 

x 

2 160 
ui 

cc 120 
m 

uj 80 

z 
z 
- 40 





KEY 
+ Unstable 
• Stable 



1 



J_ 



l 1 l 1 I 1 



4 6 8 10 12 
INNERBEDS 



14 



18 



FIGURE 2.— Number of innerbeds versus innerburden 
thickness for stable and unstable lower seam conditions. 
Adapted from Haycocks (3). 



o 

a. 



3 

cc 

I- 
X 
UJ 



< 

UJ 
CO 

cc 

UJ 
Q. 
Q- 

z> 



90 


1 1 1 

KEY 
- + Unstable 


1 




• Stable 

+ + 


/. 




/ 




8U 


' + + ♦ / 






♦/ 


• 


70 


/. 


• - 


60 


- /.. 


- 


RD 


1 1 u 


• 
1 



60 70 80 90 100 

LOWER SEAM EXTRACTION, pet 

FIGURE 3.— Relationship between percent extraction in the 
upper and lower seams and lower seam stability. Adapted 
from Ehgartner (7). 



46 



Figure 1 establishes the innerburden 
limits for interaction problems due to 
pillar loading onto an underlying seam 
in terms of the percentage sandstone in 
the innerburden. Subsequent photoelas- 
tic analysis of this pillar load transfer 
phenomenon indicates that the degree 
of layering in the innerburden is as 



important as the modulus, as shown in 
figure 2 (_7)» It was found that the rel- 
ative extraction ratios in two adjacent 
seams are extremely important in deter- 
mining whether or not interaction will be 
experienced. Higher upper seam extrac- 
tion will increase the chance of unstable 
conditions in an adjacent seam (fig. 3). 



STATISTICAL ANALYSIS 



Since the field data collected are nu- 
merous and very complex, multivariate 
statistical procedures were used to gain 
further insight into the field data. 
Statistical analyses were also used to 
derive relationships between geological 
factors, which are generally difficult to 
incorporate into other analytic models 
(_3_, 8-10). For the proposed analysis, 
variable selection was primarily deter- 
mined by former research experience and 
the availability of data. The appendix 
shows a total of 24 variables selected 
and used. Certain qualitative variables 
are used for classification purposes, and 
some of them were found very useful 
in separating the local problems due to 
in situ geological disturbances from in- 
teraction problems caused by mining exca- 
vations in adjacent seams. 

Five multivariate analysis techniques 
were used in the analyses (11-12). Ex- 
cept for regression analysis, these pro- 
cedures are similar in that they all try 
to transfer the original data set into a 



So 



Q- - I 



-2- 

-3.0-2.5-2.0 -1.5 -1.0-0.5 0.0 0.5 1.0 1.5 2.0 

PRINCIPAL 3 

FIGURE 4.— Results from principal component analysis 
(overmlnlng). 



1 1 1 1 


i 




1 1 


1 


1 s 


- 










y - 


KEY 










'o 


• Interaction 






• 


/ 




o No interaction 






• / 




- 


— 






/o 

y 




- 


- • 


• 


y 


o 






— • 


/ 


o 


o 




— 


> 










- 


/ 






o 






y 












y 












< 






o 




o 


- y 










- 


y i i ii 


i 




1 1 


°l 


1 



new data set. By data transformation, 
some specific goals can be achieved which 
may lead to better understanding of the 
original problem (13-14). 

RESULTS FROM PRINCIPAL COMPONENT 
ANALYSIS AND FACTOR ANALYSIS 

A principal component analysis was per- 
formed on eight quantitative variables. 
By Kaiser's rule (eigenvalue > 1), or 
proportional criterion, three principal 
components were extracted that explain 71 
pet of the total variance of the original 
data. Here, the cases with interaction 
can be better separated from those with- 
out interaction (fig. 4). Also, when the 
first component is large or the third 
component is small, the chance of inter- 
action problems increases. The first 
principal component has relatively higher 
loadings on variables such as COVER, UST, 
USEP, and IBT while the third component 
has higher loadings on the variables IBSP 
and LSEP. (See appendix 1 for definition 
of variables.) This indicates that the 
interaction problem may be attributed to 
cover depth, upper seam thickness, ex- 
traction on both seams, and nature of 
innerburden. It is noted that innerbur- 
den (IBT) is negatively related to the 
first component. To decrease the score 
of the first component, and thus reduce 
the possibility of having interaction 
problems, the spacing between two seams 
must increase. The results from factor 
analysis showed a similar grouping of 
variables. 

RESULTS FROM DISCRIMINANT ANALYSIS 
AND CLUSTER ANALYSIS 

Discriminant analysis was applied to 
classify the multiple-seam mines into 



47 



two groups, one with interaction problems 
and the other without, based on observed 
geological and mining conditions. Al- 
though no general way of classifying the 
observations for all mining methods was 
found, the classification of mines using 
overmining methods was successful. Fig- 
ure 5 shows the scatter plot with two 
canonical scores of each observation as 
its coordinates. It can be seen that the 
group with interaction problems is well 
separated from the group without inter- 
action effects. This implies that for 
overmining mines it is possible to pre- 
dict interaction using the discriminant 
analysis. However, severity of inter- 
action cannot be estimated by such a 
classification scheme. 

When cluster analysis, like discrimi- 
nate analysis, was used in studying all 
case study data, no natural groups or 
clusters could be identified. However, 
when this procedure was independently 
used for data from mines using overmining 
or undermining methods, it was possible 
to classify the observations according to 
location of interaction problems. 

The location where the interaction oc- 
curred was originally classified into 
five categories: below or above remnant 
pillar or pillars (A or B), below or 
above the Interface between gob area and 
large section of solid coal (C or D), or 
not as above (E). The resulted cluster- 
ing indicates that most interaction prob- 
lems can be grouped into four types of 
locations (A-D). These results suggest 





4 




1 


•1 

• 


1 1 


A 1 1 
/o 




2 




• 




• 

•• 


/ KEY 


_ 










• Interaction 


_l 






"~- — 


. 




o No interaction 


< 













— 


o 










o 




z 












~ 


o 

z 


-2 








o 


o 
o 


< 






o 






o o - 




-4 






o 




- 




_c 




1 


1 


1 1 


1 1 1 



-2.0 -1.5 -1.0-0.5 0.0 0.5 1.0 1.5 2.0 2.5 

CANONICAL 2 

FIGURE 5.— Results from canonical discriminant analysis 
(overmining). 



that the initial classification is cor- 
rect and, in addition, the interaction 
problems occurring at the same location 
may have similar geological and mining 
conditions. 

In summary, of all reported severe 
problems, 59 pet occurred either below or 
above a remnant pillar or pillars (A and 
B type locations) and 22 pet occurred be- 
low or above the interface between a gob 
area and large section of solid coal (C 
and D type locations). This demonstrates 
the possibility of interaction due to 
stress anomalies caused by isolated pil- 
lars left on the upper or lower seams. 

RESULTS FROM REGRESSION ANALYSIS 

When data were sorted into different 
mining methods, and innerburden thickness 
between adjacent seams was limited to 
less than or equal to 300 ft, good linear 
models were obtained for close-seam mines 
using stepwise regression procedures. 
For example, for mines with close seams 
and using overmining methods, the follow- 
ing regression equation can be derived: 

INTERP = -2.13 + 0.32USEP - 0.38LSEP 

+ 0.121BSP + 0.58LST. 

This model can explain 78 pet of total 
variance (R-square = 0.7799) and has a Cp 
value of 2.76, which is the best compared 
to other possible models. The error mean 
square of this model is also very small. 
In a physical sense, the degree of damage 
due to interaction in overmining can be 
related to, according to sequence of im- 
portance, upper and lower seam extraction 
percentages, Innerburden sandstone or 
strong rock percentage, and lower seam 
thickness. 

Through similar procedures, a model was 
developed for mines where the overlying 
seam has been previously mined. The re- 
gression model is given as follows: 

INTERP = 7.68 - 0.021BT - 0.06LSEP 

+ 0.27LST. 

This is also a statistically signifi- 
cant model (F-value = 8.01) and has less 



48 



variance inflation (p = 1.12). These 
models demonstrate that interaction- 
controlling factors for mines using over- 
raining methods are different from 
those using undermining methods. In this 



latter model, the important variables are 
innerburden thickness, lower seam extrac- 
tion percentage, and lower seam thickness 
(15). 



MODEL ANALYSIS 



To extend the information gained from 
case studies, analysis of various inter- 
action phenomena has been carried out us- 
ing body-loaded photoelastic and finite 
element models. These two modeling meth- 
ods are concentrated in stress analysis 
in the innerburden for undermining 
conditions. 

PHOTOELASTIC STRESS ANALYSIS 

Photoelastic models were used to eval- 
uate stress phenomena in close proximity 
to the excavation since they could best 
incorporate the effects of layering and 
bed separation, which is extremely diffi- 
cult to accomplish using numerical meth- 
ods. Peng (2^) predicted that the load 
approaches the normal background value at 
a distance of approximately four times 
the pillar width below the floor. Eh- 
gartner (7_) found that the distance 
stress was transferred depended on the 
nature of the rock below the pillar. In 
particular, he found that low-modulus 
stratified materials tended to increase 
the distance through which stress is 
transferred, while stiff isotropic mate- 
rials had the opposite effect. Results 
of this model showed that the zero-influ- 
ence bulb could extend as deep as eight 
times the pillar width in a highly strat- 
ified material (fig. 6). 

FINITE ELEMENT ANALYSIS 

Finite element analysis was used to 
evaluate the influence of thickness and 
location of sandstone layers in the in- 
nerburden and overburden on the size and 
shape of the pressure arch around the 
lower opening. This information was, in 
turn, used to locate the area in which 
the upper seam arch intersected the lower 
workings. Hudock (16), modeling in Iso- 
tropic overburden, found that the modulus 
of the material had little influence on 



arch dimensions or abutment pressures. A 
high-modulus layer above an excavation, 
however, can affect subsidence throughout 
the formation (2^, 17). The effects of 
opening geometry and material properties 
on the stability of interactive excava- 
tions in stratified rock were also eval- 
uated using finite element methods. The 
influence of pillar location, Young's 
modulus, and Poisson's ratio were deter- 
mined. This system was modeled with 
four-node isoparametric, quadrilateral, 
linear elastic, plane strain elements. 
The standard finite element routine was 
modified to include the calculation of 
the factor of safety at the center of 
each element based on the Mohr-Coulomb 
failure criterion as outlined by Wang 
(18) and Haycocks (3). The results from 
this work showed that increasing Pois- 
son's ratio resulted in improved stabil- 
ity in the roof of both upper and lower 
openings for all innerburden spacings. 
The shear stress at the pillar edge in 
the roof of the lower opening and the 
floor of the upper opening is also 



0.8 



cr 
o 

I- 
o 

& 

hi 
O 

Z 
Ul 

_l 



in 

<r 



.6- 



.4- 



.2 - 







\\ ' 


i i i i i 


\v 


KEY 




* a Thin layering 

o o Thick layering 


\\ 




- «^ 


V \ 

\Ox - 



I 2 3 4 5 6 7 8 
DISTANCE BELOW PILLAR, X W p 



FIGURE 6.— Effect of the degree of layering on the size and 
shape of the pressure bulb based on photoelastic models. 
Adapted from Ehgartner (7). 



49 



decreased, indicating less chance of mas- 
sive inner seam shear. Alternatively, 
increasing Young's modulus raised the 
tensile stress in the roof above both 
openings but decreased the maximum com- 
pressive stress in the pillars of both 
seams. The shear stress in the roof of 
the lower opening and the floor of the 
upper opening is slightly decreased in 
this condition. 

Although the finite element modeling 
method has its limitation due to the 



idealization of true conditions, it was 
felt that this method was useful in dem- 
onstrating trends as to the influence of 
variables on opening stability, even 
though the calculated stresses and factor 
of safety were only valid for comparison 
purposes. The incorporation of these 
model studies with field-observed data 
can provide realistic design criteria and 
methods for evaluating a wide variety of 
mining geometries and geologic settings. 



ANALYSIS OF UPPER SEAM SUBSIDENCE 



When dealing with interaction problems 
occurring during overmining operation, a 
different approach is used. Empirical 
methods were used to classify the upper 
seam structure and to predict upper seam 
stability. 

METHODS OF PREDICTION 

Based on statistical analysis, three 
methods have been proposed for evaluating 
upper seam stability, depending upon the 
application needs and type of input data 
available. The first method calculates a 
damage factor (_8) ; the second, or M-index 
method, uses the critical innerburden 
thickness (19); and the third, which ap- 
parently holds the most promise, is the 
subsidence factor method. 

Plots of field data show that the over- 
all damage rating can be related to the 
maximum subsidence (_8). Figure 7 shows 
the relationships between damage factor 
and maximum predicted subsidence at the 
level of the upper seam for the various 
roof conditions. Roof structural condi- 
tions can be evaluated using the various 
geomechanics classification systems, tak- 
ing into account factors that are pecu- 
liar to multiple-seam operations. An in- 
dex can be assigned to each class of roof 
as follows: 

1 = Very good roof. 

2 = Good roof. 

3 = Fair roof. 



4 = Poor roof. 

5 = Very poor roof. 

The damage rating can then be written as 

DF = ai + biS max + anxi + ai 2 x 2 

+ a^xi* + a 15 x 5 , 

where DF is damage factor and xj , j 
= 1,2,3,4,5 are indicator variables rep- 
resenting the type of roof and are de- 
fined as follows: 

xj = 1 for type j roof, 

Xj = for all others. 



o 4 



< 

< 



i r~ 

KEY 

90-pct confidence 
interval 



Roof 




0.5 1.0 1.5 2.0 

MAXIMUM SUBSIDENCE, ft 



FIGURE 7.— Nomogram for predicting upper seam stability. 
Adapted from Webster (8). 



50 



Precise determination of the exact 
mathematical relationships depends on the 
geological and spatial factors. 

STABILITY OF UPPER SEAM STRUCTURES 

Of all possible disturbances that can 
be caused to the upper seam, the most 
frequently encountered difficulty has 
been damage to the roof. Although pillar 
load increases when mining crosses the 
gob line to the solid side of the lower 
seam, as recorded by Chekan (20) , the in- 
crease is usually not large enough to 
cause any major problems. One exception 
is the situation in which pillars in the 
upper seam are located over remnant pil- 
lars of substantial size left in the 
lower seam, which transfers high stresses 
through the innerburden to the upper 
seam. 

COMPUTERIZED EVALUATION 



Roof conditions in the upper seam, on 
the other hand, are greatly affected by 
the subsidence of the strata. This is 
because the roof is most prone to tension 
created by the subsidence of the strata. 
Therefore, the location and orientation 
of the roof structures in relation to the 
subsidence trough is a very important 
factor in the design of the upper seam. 
The two most consistent places for seri- 
ous floor disturbances in an upper seam 
are over isolated pillars or small groups 
of pillars left in the underlying seam, 
and in areas above the line between 
completely mined areas, or groaves, 
and solid coal in the underlying seam 
(1_). The mechanism is the transfer of 
stresses, as in the case of remnant 
pillars. 



OF CLOSE SEAM INTERACTION 



Because of the complexities and numer- 
ous steps necessary in multiple-seam sta- 
bility evaluation and design, the process 
was incorporated into software pack- 
ages available for use on a personal com- 
puter. The aim of these programs is to 
provide engineers with means of determin- 
ing whether any difficulties will be en- 
countered and of what type they may be 
when mining in a specific condition. 

PROGRAM DEVELOPMENT 

The computer program USEAM is used for 
undermining conditions and is organized 
into two modules: the input module and 
the calculation module. The input rou- 
tine consists of a series of menus and an 
editor for entering or changing data val- 
ues. This module also provides for stor- 
age of the data on diskette for later 
modification and use. 

The calculation module uses the data 
supplied by the input routine to predict 
when strata control problems may be ex- 
pected and to suggest design parameters. 
The calculation module is further divided 
into a series of subroutines, each per- 
forming a specific calculation or set of 
calculations. The operation of this mod- 
ule is outlined in figure 8. 





Option 


























Remnant pillar 




Solid gob interface 












Pillar stress 




Abutment location 
and stress 












Lower seam stress 




Lower seam stress 


























Pillar size 














Roof stability 














Floor stability 














Output 





FIGURE 8.— Macro flowchart for the computational module 
of USEAM. Adapted from Grenoble (21-22). 



51 



Stresses for locations under a remnant 
pillar and abutment stress field are 
first calculated. The results are then 
used in evaluating the stability of the 
lower seam roof and floor, and for pillar 
designs. Output of the results is pre- 
sented in a series of tables and design 
curves. Typical input to and output from 
the program are summarized in figures 9 
and 10. 

VALIDATION OF THE PROGRAM 

The program USEAM has been validated by 
some field study data. For example, in 
the case reported by Stemple (1_) , mining 
operation was conducted under an arch 



abutment in the upper seam. The overbur- 
den in the problem area was approximately 
1,500 ft, while the innerburden consisted 
of a massive 35-ft-thick sandstone layer. 
The excess weight created by the pressure 
arch abutment caused the pillars in the 
lower seam to bump and eventually crush. 

The program predicted that the average 
stress on the lower seam will approach 
2,500 psi. From this information pillar 
dimensions on the order of 130 ft square 
are suggested for a safety factor of 
1.5. Since the actual pillars were much 
smaller than this (25 by 80 ft), it is 
obvious that they could not be expected 
to support the load. 



CONCLUSIONS 



A comprehensive study has been con- 
ducted on the problems of interactions in 
multiple-seam operations for overmining, 
undermining, and simultaneous mining, in- 
tegrating results from field studies, 
statistical analysis, and numerical and 
physical modeling. Important findings of 
this research effort are summarized 
below. 

The majority of the interaction prob- 
lems occurred in one of three modes: min- 
ning under remnant pillars, mining under 
gob line interface, and raining over gob 
line interface. 

The most important controlling factors 
are extraction percentages, innerburden 
thickness, lithological structure, and 
time delays between operations in the two 
seams . 

Roof structures can be characterized by 
taking into account the many geological 
factors and mining factors associated 
with the two adjacent seams and their in- 
nerburden, maximum subsidence, and rela- 
tive location and orientation of seam 
workings. 

Four interaction mechanisms were iden- 
tified: pillar load transfer, innerbur- 
den shearing, arching effects, and upper 
seam subsidence. The controlling mecha- 
nism may not be the same for different 
mining sequences, and it may be appropri- 
ate to consider a different set of con- 
trolling factors when a particular mining 
sequence is selected. 



Simultaneous mining is likely to cause 
more interaction problems than other 
methods and therefore should be avoided 
if possible. 

For undermining, design guidelines have 
been proposed for the various mining 
sequences and design criteria have been 
assembled on a microcomputer in a user 
friendly, interactive package. The pro- 
gram has been so written that installa- 
tion is easy. Driven by an interactive 
menu, data input and sensitivity analysis 
can be done at the fingertips. Most of 
the relevant input data can be used but 
are not strictly required to carry out a 
comprehensive analysis. In addition, a 
regression model has also been developed 
that predicts a damage factor. 

To evaluate upper seam stability in 
overmining, three methods have been pro- 
posed for the Appalachian coalfields. 
These are the damage factor method, the 
M-index method, and the subsidence factor 
method. Any one of these methods can be 
used to predict potential hazards in the 
upper seam, depending upon the applica- 
tions and type of available data. Nomo- 
grams have been prepared for quick acces- 
sibility for the field engineers. 

Ground control in multiple-seam mining 
is a complex problem; however, under most 
conditions, interaction effects and their 
magnitude can be predicted with accuracy, 
provided sufficient information is avail- 
able on the many geologic and raining 



52 



INPUT 
Mining site: Lee County, VA 

Mine under: 1. Remnant pillar 

2. Solid-gob interface 

3. Pillars in place 

Enter number: 1 

UPPER SEAM DATA 

Seam name: Pocahontas No. 5 

Depth of overburden (ft): 510 

Density of overburden (lb/cu ft): 140 

Mining height (in): 57 

Residual pillar dimensions (length and width, ft): 150,245 

Coal K factor: 0.27 

Roof character (10=excellent, l=poor): 8 

Floor character (10=excel lent, l=poor): 6 

Jointing present? (y/n) y 

Joint set No. Spacing (ft) Orientation (deg) Dip (deg) 

1 26 170 85 

2 40 80 70 

INNERBURDEN DATA 

Average thickness (ft): 55 
Total number of beds: 8 

Sandstone present? (y/n) y 







Distance above 


Bed No. 


Thickness 


lower seam (ft) 


1 


10 


30 


2 


4 


39 



Jointing present? (y/n) y 

Joint set No. Spacing (ft) Orientation (deg) Dip (deg) 
1 40 80 70 

LOWER SEAM DATA 

Seam name: Pocahontas No. 3 

Seam height (in): 48 

Roof rock characterization (10=excellent, l=poor): 4 

Floor rock characterization (10=excel lent, l=poor): 9 

Coal K factor: 0.23 
Jointing present? (y/n) n 

FIGURE 9.— Typical input data for USEAM. 



53 



OUTPUT 

Mining site: Lee County, VA 
Seam analyzed: Pocahontas No. 3 

Optimum pillar dimensions: 50 ft long 

55 ft wide 
70-ft centers 

Optimum room width: Entries 20 ft 

Crosscuts 15 ft 

Projected extraction: 51 pet 

Maximum pillar load: 4,873 psi 
Average pillar load: 1,763 psi 

Roof index (10=excel lent, l=caved): Worst 4 

Average 7 

Floor index (10=excellent, l=serious heave): Worst 5 

Average 8 

Most likely failure mechanism: Roof shear 

Potential for cracking to upper seam (0=none, 5=certain): 3 

Probability for success: Good 

Graphics options: Enter number 

1. Plot of upper seam geometry 

2. Plot of lower seam geometry 

3. Plot of superimposed geometries 

4. Contour of pillar load 

5. Contour of roof index 

6. Contour of floor index 

7. Plot of projected failure zone in cross section 

FIGURE 10.— Typical output from USEAM. 



54 



factors. Integration of the field stud- 
ies, statistical analysis, numerical and 
physical modeling, and the help of mod- 
ern computers will bring a better 



understanding of the mechanisms governing 
interaction problems and provide guide- 
lines for ameliorating negative effects. 



REFERENCES 



1. Stemple, D. T. A Study of Problems 
Encountered in Multiple-Seam Mining in 
the Eastern United States. M.S. Thesis, 
VA Polytech. Inst. and State Univ. , 
Blacksburg, VA, July 1956, 125 pp. 

2. Peng, S. S., and U. Chandra. Get- 
ting the Most From Multiple-Seam Re- 
serves. Coal Min. & Process. , v. 17, 
Nov. 1980, pp. 78-84. 

3. Haycocks, C. , M. Karmis, E. Barko, 
J. Carman, B. Ehgartner, S. Hudock, and 
S. Webster. Ground Control Mechanisms in 
Multi-Seam Mining (grant G1115511, VA, 
Polytech. Inst.). BuMines OFR 7-84, Oct. 
1983, 328 pp. 

4. Su, W. H. , S. S. Peng, and S. M. 
Hsiung. Optimum Mining Plan for Multiple 
Seam Mining. Interim Report. Dep. Min. 
Eng. , Coll. Miner, and Energy Resour. , WV 
Univ., Morgantown, WV, Aug. 1984, 71 pp. 

5. Hasler, H. H. Simultaneous vs. 
Consecutive Working at Coal Beds. Trans. 
Soc. Min. Eng. AIME, Mining Engineering, 
v. 3, No. 5, May 1951, pp. 436-440. 

6. Haycocks, C. , B. Ehgartner, M. Kar- 
mis, and E. Topuz. Pillar Load Transfer 
Mechanisms in Multi-Seam 
Min. Eng. AIME preprint 
11 pp. 

7. Ehgartner, B. L. 
Transfer Mechanisms in Multi-Seam Mining. 
M.S. Thesis, VA Polytech. Inst, and State 
Univ., Blacksburg, VA, May 1982, 164 pp. 

8. Webster, S. , C. Haycocks, and 
M. Karmis. Subsidence Interaction Ef- 
fects in Multi-Seam Mining. Paper in 
Proceedings of 2nd International Confer- 
ence on Stability in Mining. AIME, 1984, 
pp. 589-604. 

9. Dunham, R. K. , and R. L. Stace. 
Interaction Problems in Multi-Seam-Min- 
ing. Paper in Proceedings of 19th U.S. 
Symposium on Rock Mechanics. Univ. NV, 
v. 1, 1978, pp. 174-179. 



Mining. 
82-69, 



Soc. 
1982, 



Pillar Load 



10. Engineers International. Inves- 
tigation of Problems and Benefits of 
Underground Multiple Seam Coal Mining. 
Final Technical Report. U.S. Dep. Energy 
contract DE-ACO1-70ET14242, Oct. 1981, 
292 pp. 

11. SAS Institute Inc. SAS User's 
Guide: Basics. Cary, NC, 1982, 923 pp. 

12. . SAS User's Guide: Statis- 
tics. Cary, NC, 1982, 584 pp. 

13. Dillon, W. Multivariate Analysis 
Methods and Applications. Wiley, 1984, 
587 pp. 

14. Kleinbaum, D. Applied Regression 
Analysis and Other Multivariate Methods. 
Duxbury Press, Boston, MA, 1978, 556 pp. 

15. Wu, W. , and C. Haycocks. Statis- 
tical Analysis of Interaction Problems in 
Close-Proximity Multi-Seam Mines. Paper 
in Proceedings of 27th U.S. Symposium on 
Rock Mechanics. University of Alabama, 
Tuscaloosa, AL, 1986, pp. 317-323. 

16. Hudock, S. D. The Effects of the 
Pressure Arch Upon Multiple Seam Mining. 
M.S. Thesis, VA Polytech. Inst, and State 
Univ., Blacksburg, VA, Aug. 1983, 197 pp. 

17. Randolph, B. S. The Theory of the 
Arch in Mining. Colliery Eng. , v. 35, 
1915, pp. 427-429. 

18. Wang, F. D., L. A. Panek, and 
M. C. Sun. Stability Analysis of Under- 
ground Openings Using a Coulomb Failure 
Criterion. Trans. Soc. Min. Eng. AIME, 
v. 250, Dec. 1971, pp. 317-321. 

19. Zhou, Y. , and C. Haycocks. De- 
signing for Upper Seam Stability in Mul- 
tiple Seam Mining. Paper in Proceedings 
of 5th Conference on Ground Control in 
Mining. WV Univ., Morgantown, WV, 1986, 
pp. 206-212. 

20. Chekan, G. J. >} R. J. Matetic, and 
J. A. Galek. A Case Study of Ground Con- 
trol Problems Related to Multiple Seam 
Mining in the Pittsburgh and Sewickley 



55 



Coalbeds. Soc. Min. Eng. AIME Preprint 
85-325, 1985, 9 pp. 

21. Grenoble, A., C. Haycocks, and 
W. Wu. Computerized Evaluation of Near 
Seam Interaction. Paper in Proceedings 
of 2nd Annual Conference on the Applica- 
tion of Computers to the Coal Industry. 



Soc. Min. Eng., Littleton, CO, 1985, 
pp. 317-322. 

22. Grenoble, A., and C. Haycocks. 
Design Factors in Near Seam Interaction. 
Paper in Proceedings of 4th Conference on 
Ground Control in Mining. WV Univ. , Mor- 
gantown, WV, 1985, pp. 166-177. 



56 



1. 


COVER 


2. 


UST 


3. 


LST 


4. 


USEP 


5. 


LSEP 


6. 


IBT 


7. 


IBSP 


8. 


IBNL 


9. 


TIME 


10. 


INTERP 



APPENDIX. —VARIABLES SELECTED FOR DATA COLLECTION 

Quantitative Variables 

Cover depth, ft. 

Upper seam thickness, ft. 

Lower seam thickness, ft. 

Upper seam extraction percentage. 

Lower seam extraction percentage. 

Innerburden thickness, ft. 

Innerburden sandstone percentage. 

Number of layers in the innerburden. 

Delay between operations in two seams, yr. 

Degree of damage due to interaction: 

INTERP = - no damge. 

INTERP = 5 - serious damage. 

Qualitative Variables 

Location of mine. 

Names of two seams. 

Cover rock type. 

Immediate roof for upper seam. 

Immediate roof for lower seam. 

Immediate floor for upper seam. 

Immediate floor for lower seam. 

Locations and damages in upper seam. 

Locations and damages in lower seam. 

Innerburden rock type. 

Existence of water problems. 

Mining sequence. 

Surface subsidence problem. 

Location of interaction problems. 



11. 


LOCATION 


12. 


BEDNAME 


13. 


CRT 


14. 


USIR 


15. 


LSIR 


16. 


USIF 


17. 


LSIF 


18. 


USLP 


19. 


LSLP 


20. 


IBRT 


21. 


WP 


22. 


MM 


23. 


SSUBP 


24. 


LITERP 



57 



THE BUREAU OF MINES SUBSIDENCE RESEARCH PROGRAM 
By Michael A. Trevits, 1 Roger L. King, 2 and Bradley V. Johnson 3 



ABSTRACT 



The Bureau of Mines, through its Subsi- 
dence Research Program, is focusing on 
providing the mine operator with the 
ability to predict surface movements and 
effects on ground water as a function of 
mining method and geologic context. The 
program is designed for coal basins where 
high mining activity may impact land 
use requirements. In the long term, all 
coal basins and mining methods will be 
addressed. Data sets from several sub- 
sidence monitoring sites have been or 
are being collected. Data sets are now 



available from the Eastern, Interior, and 
Rocky Mountain Coal Provinces for full- 
extraction mining methods (longwall and/ 
or room-and-pillar retreat mining). At 
select sites, shallow-aquifer monitoring 
wells have also been installed to observe 
the effects of subsidence on the ground 
water system. To date, an empirical mod- 
el for subsidence prediction has been 
generated for the Northern Appalachian 
Coal Region. This paper discusses the 
status of the Bureau of Mines Subsidence 
Research Program. 



INTRODUCTION 



The involvement of the Bureau of Mines, 
U.S. Department of the Interior, in 
mining-related subsidence probably dates 
back to the Bureau's inception in 1910, 
when a mining engineer was assigned to a 
study of mine filling in the northern 
Anthracite Field of Pennsylvania (_0. 4 
The purpose of this work was to reduce 
the amount of mine waste resulting from 
imperfect mining methods, create a safe 
work environment, and reduce the settle- 
ment of the overlying ground surface. 
Results of the investigation were subse- 
quently reported by Griffith and Conner 
(2). 

Perhaps the longest sustained subsi- 
dence investigation ever made in the 
United States involved a cooperative ef- 
fort between the Bureau, the Illinois 
State Geological Survey, and the Univer- 
sity of Illinois during 1916-24. Reports 
covering the results of these studies 

'Geologist, Pittsburgh Research Center, 
Bureau of Mines, Pittsburgh, PA. 

2 Research supervisor, Pittsburgh Re- 
search Center. 

3 Staff engineer, Bureau of Mines, Wash- 
ington, DC. 

^Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this paper. 



became classics of mining subsidence 
literature. 

Between 1926 and 1932, subsidence stud- 
ies were pursued under various sections 
of the Bureau's Mining Division: Ground 
Movement Investigations Section, Subsi- 
dence Section, and Subsidence and Ground 
Movement Section. During this period, 
subsidence was studied in connection with 
iron mining in Alabama and Oklahoma, and 
iron and copper mining in Michigan. 

Although subsidence research did not 
constitute a distinct organizational ele- 
ment in subsequent years, concepts and 
data fundamental to an understanding of 
the mechanics of subsidence were devel- 
oped in the course of many other Bureau 
investigations. Included among these 
were mine roof control, pillar extrac- 
tion, longwall and caving methods of min- 
ing, physical properties of mine rocks, 
in situ stress measurements, and physi- 
cal and mathematical modeling of mine 
structures. 

The Bureau's subsidence research was 
focused in the Minerals Environmental 
Technology Program in 1978. In 1982 cer- 
tain activities of this program were 
melded with work in mining technology to 
form the current Mining Technology Pro- 
gram under which subsidence research is 
presently conducted. 



58 



THE PROCESS OF SUBSIDENCE 



During the subsidence process, the 
ground surface is changed as a result of 
the excavation of underlying coal. The 
final surface profile is determined by a 
multitude of components: thickness of 
the coal extracted, width of the opening, 
depth of coal, lithology of the strata 
between the coal and the surface, mine 
design, mining method, and other factors. 
An analysis of the typical subsidence 
profile reveals that almost every point 
of the surface will be displaced both 
vertically and horizontally, causing 
zones of tensile and compressive strain 
along the surface. These can result in 
damage to surface structures. 



One must also consider the effect on 
ground water movement in the vicinity of 
the affected area. Historically, mining 
operations have been accused of causing 
water wells to become dry and/or pol- 
luted. However, such local effects may 
be minor compared with the potential 
problems that can occur in the Western 
and Midwestern United States. Western 
underground aquifers are a major source 
of water for populated areas; in the Mid- 
west, surface drainage patterns are im- 
portant to agriculture. It is, there- 
fore, easily understood how changes in 
the ground water system could have a 
major impact. 



SUBSIDENCE MODELS 



Three types of models have gained ac- 
ceptance by subsidence researchers — phys- 
ical, analytical, and empirical. Physi- 
cal models have been used in an effort to 
develop basic understandings of the sub- 
sidence process. Examples of physical 
models include centrifuges, sand models, 
and scale models of the overburden. Var- 
ious researchers have claimed different 
findings, but the drawbacks to such mod- 
els (accuracy and expense) have precluded 
widespread acceptance of this approach. 
However, basic insight into the subsi- 
dence process may be obtained from physi- 
cal models. 

With the advent of computers , analyti- 
cal models have become more practical to 
use in subsidence research. This tool, 
in conjunction with sophisticated soft- 
ware (finite-element programs), permitted 
researchers to analyze many more cases 
in a given time than they could using 
physical modeling. The major drawback to 
analytical models is that the researcher 
must mathematically describe a physical 
process for which very little is known. 
Another major disadvantage is that 
the analytical model requires detailed 



geotechnical information about the over- 
burden that may not be readily available 
to accurately predict the subsidence pro- 
file. Such unknowns result in predic- 
tions that do not give the required de- 
gree of accuracy. 

The empirical model has gained the most 
worldwide acceptance as the approach for 
subsidence prediction. Empirical models 
are based upon the fact that given enough 
field data (final subsidence profiles), a 
mathematical function can be written to 
describe a characteristic profile. This 
approach, of course, assumes that within 
a given area the important functional 
parameters (angle of draw, subsidence 
factor, etc.) are reasonably constant. 
If this does not happen to be the case, 
adjustments to the empirical model must 
be made to make it more or less univer- 
sal. The major drawback to empirical 
models is that they require a substantial 
amount of field data before an accurate 
mathematical function can be derived. 
However, for the mine operator, this ap- 
proach should be the easiest to utilize 
in predicting mine subsidence. 



STATUS OF BUREAU OF MINES RESEARCH 



The short-term strategy is to focus the 
research program into the coal basins 
where high mining activity is coupled 



with present land use requirements. 
These areas are delineated via consul- 
tations with industry and the Office of 



59 



Surface Mining. In the long term, all 
coal basins and mining methods will be 
addressed. 

Data sets from several subsidence-moni- 
toring sites have been or are being col- 
lected. The purpose of these data sets 
will be to develop and/or validate the 



predictive models being proposed by Bu- 
reau researchers. Initial data sets are 
now available from the Eastern, Interior, 
and Rocky Mountain Coal Provinces for 
full-extraction methods (longwall and/or 
room-and-pillar with retreat mining). 



ALTERNATIVE SURVEY SYSTEMS 



Since the Bureau's Subsidence Program 
is concentrating on the development of 
empirical model(s), the collection of 
data sets from a variety of locations is 
required. A study at a mine site normal- 
ly requires the installation of hundreds 
of survey monuments, followed by numerous 
conventional ground surveys over a period 
of time. If these tasks are being car- 
ried out at a number of mine locations, 
the data collection scheme can become 
prohibitively expensive. To compare al- 
ternate surveying capabilities, a test 
grid was constructed at the Bureau's 
Pittsburgh Research Center, Pittsburgh, 
PA. The grid of monumemts was estab- 
lished in an area that was not being un- 
dermined to eliminate the variable of 
surface movement (_3). 

Conventional and high-technology sur- 
veying systems including an electronic 
distance meter-theodolite-level (EDM- 
theodolite), an automatic recording in- 
frared laser tacheometer, a global posi- 
tioning system satellite surveyor (GPS), 
aerial photogrammetry , and prototype in- 
ertial surveyor were deployed over the 
grid during a 1-month period. 

Highlights of the systems used follow: 

The inertial system is extremely porta- 
ble and can be placed in surface vehicles 
and aircraft. Sources of survey error 
include accelerometer measurement caused 
by thermal effects, platform drift rate 
caused by vibration variations and ther- 
mal transients, and environmental effects 
such as variations in the earth's gravi- 
tational field and temperature varia- 
tions. Survey data for the Bureau study 



could not be used owing to a computer 
malfunction. 

The GPS system, when completed, will be 
able to determine receiver position in- 
stantaneously. Constraints are the high 
amperage required to power the field 
system, sensitivity to temperature 
variations, and site selection. This 
system was used to accurately determine 
the position of the control monuments, 
but the expense and time to survey the 
grid of monuments at the Bureau's study 
site precluded its use over the entire 
grid. 

The EDM-theodolite system is portable, 
rugged, and relatively simple to operate. 
The system is sensitive to windy and/or 
rainy conditions and is subject to errors 
due to hand tabulations and computations. 
This system was used as the base for the 
Bureau's study because it is the most 
commonly used method for making subsi- 
dence measurements in the field. 

The tacheometer system is completely 
automated for rapid recording, comput- 
ing, and data generation. The system is 
subject to the same environmental con- 
straints as the theodolite system. 

The aerial photogrammetry system can be 
used for inaccessible areas. All data 
are gathered simultaneously, providing a 
permanent record that can be reevaluated. 
Surveys cannot be conducted during incle- 
ment weather, and data for control points 
must be visible on the photographs. 

The results of the Bureau's study 
showed that the three-dimensional dis- 
placements from the base (EDM-theodolite) 
were almost identical for the tacheometer 
(0.25 ft) and photogrammetry (0.24 ft). 



GROUND SURFACE AND STRUCTURE RESPONSE 



In 1985, the Bureau of Mines initiated 
a cooperative mine subsidence research 
program with the State of Illinois (4-_6)« 
The purpose of this effort is to jointly 



fund research to develop guidelines for 
underground mining methods that could 
maximize coal recovery while preserving 
farmland productivity. 



60 



In 1985, the research effort evaluated 
and selected sites at operating mines 
that best fit the program's goals. Ini- 
tial surface and subsurface studies in- 
clude characterizing the response of 
overburden to subsidence, assessing the 
impact of subsidence on ground water, 
monitoring and characterizing subsidence 
profiles, collecting available data to 
build data bases on subsidence and rock 
mechanics, testing floor materials in 
mines, and evaluating crop production in 
subsided areas. 

Because much of the farmland in Illi- 
nois is flat with water table depths of 
2 to 5 ft, differential surface ground 
movements could alter subsurface hydro- 
logical patterns and affect surface 
drainage. The Bureau of Mines is cur- 
rently studying the effects of subsidence 

EFFECTS ON 

To properly characterize the effects of 
mining-induced subsidence on the local 
hydrological system, a coordinated data 
collection scheme must be constructed. 
All monitoring points (weirs, wells, 
etc. ) should be installed in advance of 
mine development in the area of study. 
It should be noted that these data will 
represent base conditions, and enough 
preliminary data should be collected to 
allow for seasonal fluctuations and well 
stabilization if applicable. 

One such study was recently completed 
by the Bureau of Mines at a mine site in 
western Pennsylvania (_7_). Th e purpose of 
this work was to observe the effects of 
longwall mining in five water wells which 
typified local domestic water systems 
(fig. 1). Two very small streams and a 
spring located in the vicinity were also 
monitored to establish flow rate charac- 
teristics. The overburden above the coal 
unit being mined (Pittsburgh Coalbed) 
ranged from 750 to 1,000 ft thick. 

Results of the study showed that there 
was no pronounced change in the overall 
quality between premining and 1 yr after 



on surface drainage patterns to determine 
critical slope changes that may cause 
drainage problems detrimental to crop 
production. Such differential movements 
also can cause damage to structures and 
utilities, not only in Illinois but po- 
tentially in any area affected by under- 
ground mining. 

By the end of 1984, more than 1,500 
claims for structural damage (of which 25 
pet were attributed to mine subsidence) 
had been filed with the Illinois Mine 
Subsidence Insurance Fund. To better un- 
derstand how surface ground movements 
from high-extraction coal mining affect a 
residential foundation, the Bureau of 
Mines built two foundations which are 
currently being monitored for vertical, 
horizontal, and differential movements. 

HYDROLOGY 

the longwall face passed the water well 
profile (fig. 2). Water levels in the 
well located near the centerline of the 
panel began to decline when the longwall 
face had approached within 500 ft of the 
well. The water level continued to fall, 
and the well went dry about 2 months af- 
ter the face had passed beneath it, at 
which time the face was approximately 500 
ft beyond the well location. The water 
levels in the two wells located 100 and 
300 ft outside the rib line of the panel 
declined some 15 to 30 ft as a result of 
mining but recovered to near premining 
levels about 10 months after the longwall 
face passed by. The well located more 
than 500 ft outside the rib line of the 
panel showed no detectable change in flu- 
id level as a result of mining. No evi- 
dence of mining effects on the small 
streams or springs located within 1,200 
ft of the panel could be detected. 

It should be noted that the results of 
this study apply to the local conditions 
of this mine site, which may differ from 
conditions at other sites. 



SUBSIDENCE MODEL 



The specific lithology over coalbeds in 
the Northern Appalachian Coal Region, 
created by highly resistive limestone 
and sandstone units with relatively 



shallow overburden, has precluded the 

use of predictive methods developed for 

European mining conditions. Bureau of 

Mines work to date has resulted in the 



61 



1,000 

j I 




LEGEND 

(^ ) Surface contour 

Water well 

I Outline of longwall panel 
— _ J Outline of proposed panel 

▼ V- notch weir 

/• Spring 



FIGURE 1.— Surface features and longwall panel. 



development of a model that is suitable 
to the geological and mining conditions 
in the northern Appalachian area (8). 
Because of these conditions, the subsi- 
dence coefficient varies within the area 
of the subsidence, trough. The effects 
of lithology in the form of a variable 
subsidence coefficient were separated for 
each site studied by introducing a corre- 
lation between hypothetically homogeneous 
overburden and existing lithological con- 
ditions, while providing for different 
mining situations, e.g., underground 
geometry and thickness of the overburden. 



To develop the model, data from 11 
longwall test sites were used in a re- 
gression analysis. The width of the in- 
dividual panels studied ranged from 400 
to 900 ft, and the overburden thickness 
was 345 to 910 ft. For each panel, the 
characteristics of the variability of the 
subsidence coefficient along individual 
profilelines were defined. The regres- 
sion analysis of the subsidence coeffi- 
cient from all of the test sites relative 
to the edge of the panel yielded a third- 
degree polynomial equation with a corre- 
lation coefficient of 0.99. 



62 



DISTANCE OF FACE FROM 
WATER WELL SECTION, ft 



2,000 - 
3,000 



15° angle to limit of 
major surface deformation- 



/ L 



Longwall paneh 



/ 

1/ 

l// 
7/ 



I I 
I I 
I I 

// 

/ 



25° angle to limit of 

detectable surface 

deformation 



/ 



Surface 






I i i_ 



500 



Scale, ft 



V 



FIGURE 2.— Relation of water levels to longwall face advance. 



63 



Although very good results for the mod- 
el have been obtained through sensitivity 
tests, the model must be considered pre- 
liminary. The model has been constructed 
using a limited amount of field data for 
relatively similar mining conditions. 
Its applicability in other mining areas 
must be tested. 

To facilitate the use of the predictive 
model, a program was written in BASIC 
language for use on personal computers. 
The program requires approximately 48K of 
memory and has been constructed so that 
hard copies of the subsidence predictions 
are possible. The inherent beauty of the 
predictive model is that it can be oper- 
ated by individuals with little or no 
previous knowledge of the theory of sub- 
sidence, it is simple and fast in compar- 
ison with existing predictive methods, 
and it eliminates the use of inaccurately 
estimated functional parameters, e.g., 
maximum subsidence, location of the in- 
flection point, etc. necessary with ex- 
isting predictive methods (fig. 3). 




FIGURE 3.— Comparison of field data, subsidence model 
prediction and data obtained using various European 
methods. 



SUMMARY 



The Bureau's Subsidence Research Pro- 
gram has been constructed to address 
nearly every aspect of the phenomenon 
of subsidence and the results of ground 
movement. The program is currently fo- 
cused on high-coal extraction systems; 
however, in the long term all coal basins 
and mining methods will be investigated. 
Work completed to date includes prelimi- 
nary observations of the effects of long- 
wall mining on shallow water sources, 
delineation of the impact of mining on 



primary and secondary land use, and de- 
velopment of a prediction model that is 
fast and simple to use. Future research 
is scheduled to quantify the effects of 
mining on the hydrological system in a 
variety of geologic environments, surface 
areas (including structures) will be in- 
strumented to observe movement as mining 
progresses beneath, and the prediction 
model will be modified and/or updated for 
coal basins outside the Northern Appala- 
chian Coal Region. 



REFERENCES 



1. Quan, C. K. Overview of the Bureau 
of Mines Subsidence Research Program. 
Pres. at Soc. Min. Eng. AIME 1979 Annual 
Meeting (New Orleans, LA, Feb. 18-22, 
1979). Soc. Min. Eng. AIME preprint 79- 
84, 1979, 9 pp. 

2. Griffith, W. , and E. T. Conner. 
Mining Conditions Under the City of 
Scranton, Pa., Reports and Maps. BuMines 
B 25, 1912, 89 pp. 

3. Krantz , G. W. , and J. C. LaScola. 
Longwall Mine Subsidence Surveying — An 



Engineering Technology Comparison. Paper 
in Mine Subsidence Control. Proceedings: 
Bureau of Mines Technology Transfer Sem- 
inar, Pittsburgh, PA, September 19, 1985, 
comp. by Staff, Bureau of Mines. BuMines 
IC 9042, 1985, pp. 2-12. 

4. Powell, L. R. Subsidence Studies 
in the Illinois Coal Basin. BuMines Re- 
search 85, 1986, pp. 36-37. 

5. DuMontelle, P. B. IlHnois Mine 
Subsidence Research Program. IL Geol. 
Surv. , 1985, 2 pp. 



64 



6. Powell, L. R. , and P. B. DuMon- 
telle. The Illinois-Bureau of Mines Co- 
operative Mine Subsidence Research Pro- 
gram. Paper in Proceedings of the Second 
Conference on Ground Control Problems in 
the Illinois Coal Basin. Southern IL 
Univ., Carbondale, IL, 1985, pp. 13-17. 

7. Moebs, N. M. , and T. M. Barton. 
Short-Term Effects of Longwall Mining on 
Shallow Water Sources. Paper in Mine 
Subsidence Control. Proceedings: Bureau 



of Mines Technology Transfer Seminar, 
Pittsburgh, PA, September 19, 1985, corap. 
by Staff, Bureau of Mines. BuMines 
IC 9042, 1985, pp. 13-24. 

8. Adamek, V. , and P. W. Jeran. Pre- 
diction of Subsidence Over Longwall Pan- 
els in the Northern Appalachian Coal Re- 
gion. Pres. at Soc. Min. Eng. AIME 1985 
Fall Meeting (Albuquerque, NM, Oct. 16- 
18, 1985). Soc. Min. Eng. preprint 85- 
404, 1985, 16 pp. 



65 



SUBSIDENCE OVER CHAIN PILLARS 
By P. W. Jeran 1 and V. Adamek 2 



ABSTRACT 



Subsidence over two or more adjacent 
longwall panels and the intervening chain 
pillars was monitored by the Bureau of 
Mines at four mines in the Northen Appa- 
lachian Coal Basin. The magnitude of the 
subsidence over the chain pillars ranged 
from 0.06 to 1 ft. The width of the 
chain pillars affects the shape of the 
subsidence curve. Wider chain pillars 
yield a wider area of minimum subsi- 
dence. Comparison of the field-measured 



subsidence with precalculated subsidence 
over the chain pillars indicates a range 
of pillar deformation. The data show 
that at three of the sites additional 
subsidence was induced over the first 
panel by the mining of the second panel. 
Curves of the additional subsidence are 
similarly shaped for these sites. This 
indicates that with sufficient data a 
model to predict subsidence over chain 
pillars could be developed. 



INTRODUCTION 



Early monitoring of subsidence over 
room-and-pillar workings showed that some 
exhibited fairly regular subsidence, 
while others were highly irregular. The 
irregular subsidences were thought to re- 
sult from the incomplete extraction of 
the pillars where anything from stumps to 
entire pillars were left depending on lo- 
cal mining conditions. 

When the Bureau of Mines began a proj- 
ect to develop a method of subsidence 
prediction, it was reasoned that data ob- 
tained from monitoring over longwall op- 
erations would be the least complicated 
and therefore the easiest to interpret 
the process of subsidence in this coun- 
try. Longwall mining removes all the 
coal and any observed surface subsidence 
irregularities are therefore related to 
the response of the overburden to that 
removal. With room-and-pillar mining, 
there would always be the question of 
what portion of the subsidence was due to 
overburden movement and what was caused 
by incomplete mining. 

The monitoring of selected longwall 
panels has shown that subsidence above 
the centerline is not uniform (1). The 



1 Geologist. 
2 Mining engineer. 
Pittsburgh Research Center, 
Mines, Pittsburgh, PA. 



Bureau of 



lack of uniformity may be due to local 
stratigraphic inhomogeneities or varia- 
tion in the thickness of coal extracted. 
Sufficient field data are not available, 
at this time, to resolve this question. 

A predictive method has been developed 
(2) using data from half profiles extend- 
ing from the centerline of a panel out 
over the solid coal. The method has been 
proven valid for the Northern Appalachian 
Coal Basin (3_). Using the principle of 
superposition, the subsidence over the 
chain pillars separating two longwall 
panels can be predicted by adding the 
subsidence caused by the mining of each 
of the two panels illustrated in figure 
1. This assumes that the chain pillars 
resist deformation and behave in the same 
way they would if the second panel had 
not been mined. If these assumptions are 
not met, then the actual subsidence 
should exceed the prediction. 

Chain pillars are used to maintain 
the integrity of the entries surrounding 
longwall panels, and there is wide vari- 
ation in the dimensions of chain pillars 
used in this country. Some are designed 
to yield, while others are sized to 

-^Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this paper. 



66 



resist movement. To provide additional 
support at the working place, some mine 
operators backfill the entries between 
the chain pillars. Others install crib- 
bing or posts, and in some instances they 
may do nothing. Each of these support 
materials will behave differently when it 



becomes part of the gob; hence, effects 
on subsidence will also vary. This paper 
examines the surface subsidence over the 
chain pillars between longwall panels at 
four sites. The data were obtained using 
standard surveying methods. 



DISCUSSION 



The Bureau of Mines has monitored sub- 
sidence over two or more adjacent panels 
wherever possible. Since this is a mul- 
tiyear task, great care must be taken to 
assure long-term stability of survey con- 
trol points. Any undetected movement of 
the control points can yield erroneous 
data on ground movement. Data from four 
study sites were chosen based upon data 
reliability as well as difference in 
mining geometry and measured subsidence. 
While these data sets are insufficient to 
construct a model of the subsidence oc- 
curring over chain pillars, their study 
can provide some insight to the process. 

The four study sites are located in the 
Northern Appalachian Coal Basin, south- 
western Pennsylvania and north-central 
West Virginia. At each site subsidence 
was monitored over at least two longwall 
panels and the Intervening chain pillars. 
The main criterion for site selection was 
that monitoring begin sufficiently dis- 
tant from mining to preclude any prior 
subsidence movements. All surveying was 
conducted using EDM and precision level- 
ing equipment. Survey control points 
were established and maintained through- 
out the monitoring. Initialization of 
survey points occurred prior to any pos- 
sible movement. Profiles were located 



Predicted 

subsidence 

from panel 2 



o 
| - 2 

CO 
CD 

z> 
co -3 




Predicted 
subsidence 
from panel I 



Subsidence from pane 
plus panel 2 — 




Panel I 



Panel 2 



100 200 300 400 500 600 700 800 

LOCATION, ft 

FIGURE 1.— Subsidence over chain pillars based on super- 
position. 



far enough from the beginning and end of 
each panel to eliminate any reduction in 
subsidence caused by these areas. 

The generalized stratigraphy over each 
study site as illustrated by columnar 
sections is shown in figure 2. The sec- 
tions are based on coreholes drilled at 
or very near each study site. The strata 
are typical of the Northern Appalachian 
Coal Basin, being composed of alternat- 
ing layers of varying thicknesses of 
resistant rocks (limestones and sand- 
stones) separated by shale. 

900 r 



800 



700- 



600 



«= 500 

L±J 

_l 

<-> 400 
co 



300 



200 



I00- 



L 



IT I If 






LEGEND 

EUSI Sandstone 

Limestone 

Shale 

Coal 



A B C D 

MINE 

FIGURE 2.— Generalized columnar sections. 



67 



Table 1 contains a summary of the min- 
ing geometries. The panel widths range 
from 625 to 1,000 ft. The coal extracted 
averaged 5.5 to 6 ft thick. Chain pillar 
sizes and numbers as well as overbur- 
den thicknesses varied among the sites. 
Lithologic composition of the overbur- 
den was similar in that the bulk of the 
resistant strata at each site is within 
the first 300 ft above the mine. 

Mine A is located in northern West Vir- 
ginia on the Pennsylvania border. Three 
panels were monitored at this site. Each 
was 5,300 ft long by 625 ft wide. The 
panels were developed using a four-entry 
system driven on 100-ft centers except 
for the fourth entry, which was driven 
on 60-ft centers to accommodate a yield 
pillar on the gob side of the panel. The 
average extracted height was 6 ft. The 
overburden ranged from 700 to 950 ft 
thick over the first two panels. It ex- 
ceeded 1,100 ft over the third panel. 
The lithologic composition of the over- 
burden is shown in figure 2. The first 
250 ft above the mined coalbed contains 
most of the resistant strata in the 
column. 

Mine B is located in southwestern Penn- 
sylvania. The panels were 630 ft wide. 
Panel 1 was 4,700 ft long, and panel 2 
was 5,650 ft long. The panels were de- 
veloped using four entries driven on 
90-ft centers. The total width of the 
chain pillars and entries was 280 ft. 
The overburden ranged from 785 to 890 ft 
thick. The columnar section (fig. 2) 
shows resistant strata between 30 and 190 
ft above the coalbed that was mined and a 
45-ft-thick sandstone 135 ft above that. 
An average 6 ft of coal was extracted 
from these panels. 



Mine C is located in north-central West 
Virginia. Panel 1 was 1,000 ft wide, and 
panel 2 was 950 ft wide. They were sep- 
arated by a pair of chain pillars 62 ft 
wide. Total width of the chain pillars 
and entries was 160 ft. The overburden 
ranged from 665 to 690 ft thick. Most of 
the resistant strata are between 60 and 
235 ft above the mined coalbed (fig. 2). 
The extraction thickness averaged 5.5 ft. 

Mine D is located in the panhandle of 
West Virginia between Ohio and Pennsyl- 
vania. The panels were 605 ft wide by 
3,000 ft long. They were developed using 
a four-entry system. Two of the three 
chain pillars were 85 ft wide, and one 
is 65 ft wide. The total width of the 
chain pillars and entries was 280 ft. 
The overburden ranged between 510 and 
660 ft thick; its composition is shown 
in figure 2. The first 250 ft above 
the coal contain most of the resistant 
strata. The extraction thickness aver- 
aged 5.5 ft. 

The Bureau subsidence prediction mod- 
el was used to determine the amount of 
subsidence caused by the mining of each 
panel. Applying the principle of super- 
position, the subsidence resulting from 
the mining of each panel was added to- 
gether to obtain the subsidence over the 
the chain pillars between the adjacent 
panels at each mine. Comparing these 
data to the field measurements (figs. 3- 
6) showed only mine D in close agreement 
with the model. The other mines show 
that subsidence greater than predicted by 
the model had occurred. Given that the 
model predictions are reasonable, it must 
be concluded that at three sites the 
overburden reacted as if the chain pil- 
lars had deformed. 



TABLE 1. - Mining geometries 





Overburden 

thickness, 

ft 


Panel 

width, 

ft 


Coal 

height, 

ft 


Chain pillar 


Mine 


Centerline 

to 
centerline 


Total 
width, 
ft 




680-950 

745-910 
660-710 

510-660 


625 

630 
1,000 

605 


6.0 

6.0 
5.5 

5.5 


f 2 at 100 ft 
1 1 at 60 ft 
3 at 90 ft 
2 at 80 ft 
f 2 at 100 ft 
1 1 at 80 ft 


| 260 

280 
160 

| 280 



68 



CD 

in ' 3 - 



-i r~ 



oooooooo, 



o • 



KEY 

o Prediction 
• Field 



o° Panel I Chain pillars • 



Panel 2 



® 



100 200 300 400 500 600 700 800 900 

LOCATION, ft 

FIGURE 3.— Predicted and measured subsidence over chain 
pillars at mine A. 



LU 
O 

UJ 


-1 
-2 






0000000o 

o°° 

o 
o 

o .« 

• 


°o 


i i 

o 

o 
o 

•• ° 

• 

• ° 


Q 
CO 

m 








°» KEY 
o • o Prediction 




• o 




-3 






o • • Field 

• 




•o 

• 




-<a 


©8 


« 


Panel 1 Chain pillars— ^ 




Panel 2 o 

* a 

1 1 Ij 




i ■■MBMHHM 


1 1 



100 200 300 400 500 600 700 800 900 
LOCATION, ft 

FIGURE 4.— Predicted and measured subsidence over chain 
pillars at mine B. 



C -I 

o 

S -2 
g 

V) 

CD 

co "3 



-4 



1 ' 1 - k 1 1 1 
o°°° 0t00 °o 



° •••••••• ° 

o # » • • o 

• • 
o • 8 
• • 
o • o • 

e # KEY o» " 


• o Prediction ° 
^e • Field °*. 


Ponel 1 Chain pillars^ p ane | 2 



200 300 400 500 600 700 800 900 1,000 
LOCATION, ft 

FIGURE 5.— Predicted and measured subsidence over chain 
pillars at mine C. 



LU 

o 

S -2 



1 1 — 




o i 8 8 


* 
o 


— r 
• 




1 








e • 




o 










S 








• 






- 










o 




- 




• 












8 




o 














- 






KEY 








o 









o Prediction 








• o 




• 




• Field 








• 


~ 


t Panel 1 




Chain pillars -7 




Li_ 


Panel 2 


8 n 







01 
CD 
3 
CO 



100 200 300 400 500 600 700 800 900 
LOCATION, ft 

FIGURE 6.— Predicted and measured subsidence over chain 
pillars at mine D. 



SUBSIDENCE CALCULATIONS 



69 



Subsidence can be calculated from the 
product of the extracted thickness, the 
subsidence coefficient, and the effi- 
ciency coefficient. Based on the Bureau 
model, the subsidence coefficient is con- 
stant beyond the edge of the panel. 
Therefore, the difference between the 
predicted subsidence and the field mea- 
surements is proportional to the prod- 
uct of the efficiency coefficient and a 
change in the height of the chain pil- 
lars. The efficiency coefficient can be 
calculated using an algorithm based on 
Bals' Theory (_4). The deformation of the 
chain pillars expressed as a reduction in 
thickness can be calculated from the dif- 
ferences between the predicted and mea- 
sured subsidence divided by the effi- 
ciency coefficient. Figure 7 illustrates 
the results of these calculations for 
each site. The results show a logical 
progression from no reduction in pillar 
thickness at mine D to the largest reduc- 
tion at mine C. With increasing average 
reduction in thickness of the pillars, 
the edges appear more affected than the 
center. The mechanics of this apparent 
pillar deformation are unknown. Possi- 
ble causes are the punching of the chain 
pillars into the floor rock, the crushing 
of the chain pillars, and the movement of 



the roof rock into the voids surrounding 
the pillars. 

At mines A, B, and C, where the raining 
of the second panel caused additional 
subsidence to occur over the first panel, 
the difference in subsidence between the 
two shows similar trends (figs. 8-10). 
At these sites the additional subsidence 
decreases linearly with distance from the 
chain pillars toward the first panel. 
This indicates that with sufficient data 
sets it should be possible to create a 
predictive model for subsidence over 
chain pillars. The data available at 
this time, however, do not allow the de- 
termination of the overburden thickness, 
panel width, and chain pillar configura- 
tions that will prevent the additional 
subsidence occurring over the first panel 
when the second panel is mined. 

The time factor associated with chain 
pillar stability is also unknown. Should 
the chain pillars or the cribbing or oth- 
er supports deteriorate, then additional 
subsidence should occur. Therefore, it 
appears that reducing the size of chain 
pillars in the gob would assure long-terra 
surface stability. Ideally, the elimina- 
tion of chain pillars would minimize sur- 
face deformation when the subsidence pro- 
cess is completed. 



70 



KEY 

Reduction in thickness 
across chain pillars 



ZZC 




Mine A 



Mine B 




FIGURE 7.— Reduction in height across chain pillars. 



LU 

S -2 

q 

(D 

CO 

=> -3 



! 


i 


i i 

•• 

• 
• 

• 


- 






• 
• 

• 


" 


Panel 1 

i 


Chain pillarsy 


• 

Panel 2 

i i 





200 



400 600 

LOCATION, ft 



800 



1,000 



FIGURE 8.— Difference in subsidence between mining of 
panels 1 and 2 at mine A. 




400 600 

LOCATION, ft 



800 1,000 



FIGURE 9.— Difference in subsidence between mining of 
panels 1 and 2 at mine B. 



uj -2 

Q 
CO 

m 

3 -3 



1 

Panel 1 


Chain 


i 
pillars-? 


• 
• 
• 
• 

• 

Panel 


i 

• 

• 
• 
• # 

2 


^ 





200 400 600 800 

LOCATION, ft 



1,000 1,200 



FIGURE 10.— Difference in subsidence between mining of 
panels 1 and 2 at mine C. 



71 



SUMMARY AND CONCLUSION 



Chain pillar 
left in place 
tween adjacent 
retreating of 
come part of 
subsidence is 
profile across 
rate troughs, 
over chain pil 
Northern Appal 
face movements 
ing between 0. 
troughs subsid 
The data indi 
width of the 
minimize the 
between that 
of the panels 



s are the blocks of coal 

to protect the entries be- 

longwall panels. With the 

the longwall face, they be- 

the gob. Their effect on 

to break up the subsidence 

several panels into sepa- 

Measurements of subsidence 

lars at four mines in the 

achian Coal Basin show sur- 

over chain pillars rang- 

06 and 1 ft. The adjacent 

ed between 3.5 and 3.9 ft. 

cate that minimizing the 

chain pillars would also 

difference in subsidence 

occurring over the center 

and that over the chain 



pillars. The subsidence prediction model 
developed by the Bureau predicts subsi- 
dence over chain pillars only when the 
chain pillars behave as if the second 
panel had not been mined. The similarity 
of the shapes of the curves of difference 
in subsidence between mining of first and 
second panels indicates that with suf- 
ficient data the Bureau model could be 
expanded to include the prediction of 
subsidence over chain pillars. The long- 
term stability of chain pillars is un- 
known, and the materials used to support 
the intervening entries are subject to 
eventual deterioration. Therefore the 
potential exists for further subsidence 
using the present configurations of chain 
pillars. 



REFERENCES 



1. Jeran, P. W. , and T. M. Barton. 
Comparison of the Subsidence Over Two 
Different Longwall Panels. Paper in Mine 
Subsidence Control. Proceedings of Tech- 
nology Transfer Seminar, Pittsburgh, PA, 
September 19, 1985, comp. by Staff, Bu- 
reau of Mines. BuMines IC 9042, 1985, 
pp. 25-33. 

2. Adamek, V., and P. W. Jeran. Pre- 
calculation of Subsidence Over Longwall 
Panels in the Northern Appalachian Coal 
Region. Paper in Mine Subsidence Con- 
trol. Proceedings of Technology Transfer 
Seminar, Pittsburgh, PA, September 19, 
1985, comp. by Staff, Bureau of Mines. 
BuMines IC 9042, 1985, pp. 34-56. 



3. Jeran, P. W. , V. Adamek, and M. A. 
Trevits. A Subsidence Prediction Model 
for Longwall Mine Design. Paper in 
Proceedings of Longwall USA Conference 
(Pittsburgh, PA, June 17-19, 1986). In- 
dustrial Presentations West Inc. , 1986, 
pp. 101-112. 

4. Adamek, V., and P. W. Jeran. Eval- 
uation of Existing Predictive Methods for 
Mine Subsidence in the U.S. Paper in 
Proceedings of First Conference on Ground 
Control in Mining. WV Univ. , Morgantown, 
WV, 1981, pp. 209-219. 



72 



STUDY OF DEWATERING EFFECTS AT AN UNDERGROUND LONGWALL MINE SITE 
IN THE PITTSBURGH SEAM OF THE NORTHERN APPALACHIAN COALFIELD 

By Gregory E. Tieman 1 and Henry W. Rauch 2 



ABSTRACT 



Dewatering effects from longwall raining 
were studied for a mine site in south- 
western Pennsylvania as part of a re- 
search project funded by the Energy Re- 
search Center of West Virginia University 
with contributions by the U.S. Bureau of 
Mines. The mine showed evidence of dewa- 
tered streams and ground water supplies. 
Water sources located much above base 
level (major stream level) and over or 
adjacent to recently mined longwall pan- 
els were partly to completely dewatered, 
probably owing to downward leakage along 



subsidence fractures. These lost waters 
did not migrate to the deep mine because 
of its thick overburden (at least 500 
ft), but instead flowed laterally over 
confining strata to discharge at nearby 
streams. Many affected water supplies 
recovered partially, and all streams re- 
covered fully within 1 to 3 yr following 
longwall mining. Spring and well recov- 
ery occurred most frequently near local 
stream level where newly formed springs 
were also common. 



INTRODUCTION 



BACKGROUND 

The coal produced from underground min- 
ing is an integral component of the econ- 
omies of the Northern Appalachian Coal 
Basin. Unfortunately, adverse environ- 
mental effects can be associated with 
underground coal mining. These effects 
include (1) surface damage resulting from 
land subsidence, (2) degradation of 
ground and surface water quality, and 
(3) stream and aquifer dewatering with 
associated fluctuations of the ground wa- 
ter levels. Federal regulations are in- 
tended to ensure the health and safety of 
the public and to minimize potential dam- 
age to the environment. The regulations 
require the mining permit applicant to 
"...identify the extent to which the pro- 
posed underground mining activities may 
proximately result in contamination, di- 
minution, or interruption of an under- 
ground or surface source of water within 

Hydrologist, Malcolm-Pirnie, Inc., 
White Plains, NY. 

2 Professor of geology, Department of 
Geology and Geography, West Virginia Uni- 
versity, Morgantown, WV. 



the proposed mine plan or adjacent area 
for domestic, agricultural, or other le- 
gitimate use" (_j_). 

The partial or complete loss of ground 
water supplies and streams can be of sig- 
nificant impact in rural areas of the 
Northern Appalachian Coal Basin. Al- 
though mine subsidence damage to a sur- 
face structure may be significant, the 
loss of water is often a longer term 
problem. For example, it is inconvenient 
to haul replacement water for domestic 
household use, and replacing lost water 
supplies for livestock often is unecono- 
mical. Therefore, an assessment of the 
impacts of underground coal mining on 
nearby water supplies is important to 
help avoid such impacts and to ensure the 
optimum replacement of lost supplies. 

The assessment of mining impacts upon 
local water supplies, particularly im- 
pacts on the longwall mining method, have 
been difficult to determine because of a 
general lack of documented evidence. The 

■^Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this paper. 



73 



longwall mining method is a nearly total 
coal extraction system that is account- 
ing for an increasing percentage of un- 
derground coal production in the North- 
ern Appalachian Coal Basin (2^). In this 
basin there are numerous reports of 
negatively impacted domestic water 
wells, springs, ponds, and streams, but 
there have been few scientific attempts 
to quantitatively relate these reports 
to parameters such as type of under- 
ground mining, age of mines, topography, 
and mine overburden stratigraphy and 
thickness (_1_) . 

The overall objective of this study was 
to determine the impacts of two under- 
ground coal mines on streamflow and 
ground water supply levels. One study 
site (reported on in this paper as study 
site Y) is located in Greene County in 
southwestern Pennsylvania; the second 
site (study site Z), which is not re- 
ported on in this paper, is located in 
Monongalia County in northern West Vir- 
ginia (fig. 1). Both of these under- 
ground coal mines utilize the longwall 
raining method in the Pittsburgh Coal 
Seam. These two coal mines were selected 
for study because of the cooperation of 
the mining companies involved as well as 
the similarities in geologic setting, 
mined coal seam, and mining methods used. 
Both of these mines were studied for wa- 
ter quantity impacts by mining in an M.S. 
thesis study by Tieman (3_). The terrain 
in this portion of the unglaciated Appa- 
lachian Plateau consists of rounded hills 
separated by narrow stream valleys with a 
topographic relief of 300 to 450 ft. The 
mine overburden at both sites consists of 
shale, mudstone, claystone, fire clay, 
sandstone, limestone, and coal with a 
thickness of 650 to 1,100 ft above the 
Pittsburgh Coalbed. The effects of coal 
mining on water resources in Greene Coun- 
ty, PA, have been generally characterized 
by the U.S. Geological Survey (4_~_5)« 

METHODS OF INVESTIGATION 

A domestic water well and spring inven- 
tory at mine Y yielded data on 58 water 




LEGEND 
— State boundary 
City 



100 

J 



Scale, miles 



FIGURE 1.— Location map of mine study sites Y and Z. 

wells and springs in February, June, and 
July, 1986. The inventoried water sup- 
plies were located above mine Y and with- 
in 2,000 ft of the nearest longwall or 
room-and-pillar mine area (fig. 2). The 
data collected represented the (1) static 
water table depth, (2) spring yield, 
(3) total well depth, (4) depth of well 
casing, (5) observed subsidence damage, 
and (6) hydrologic changes to the water 
source. The data collected were from 
published reports, information from mine 
company personnel and owners of the do- 
mestic water supplies, and from measure- 
ments made by the senior author. The 
depth to static water level was deter- 
mined when the well top was accessible 
by utilizing a Soiltest electric water 
level meter. Spring discharge in gal- 
lons per minute was determined with a 
2-gal graduated bucket and stopwatch. 
Stream discharge at 28 sampling points on 
8 streams was measured by the authors on 
either or both June 6 and July 24, 1986. 
Streamflow was measured during baseflow 
conditions, with no measurable precipita- 
tion during the previous 24 h so as to 
have stable and easily measurable flows. 
Stream discharge in gallons per minute 
was measured using a 6-in PVC pipe and a 
2-gal graduated bucket with a stopwatch. 
A Parshall flume with a 50-mm throat was 



74 




■ ;33 



."' -■■ :.\\ *■:>. :■»■ ' . ■»' ■ i.VJ.~:U iVl^^f'-'-TA ' - - -^ 



(a 



^ 



• 40 



'fle 



39 •„ 



c o| 



"48 



£7 4 « 7 4>45 46 '55~54 V 



"52 



36 



37 "38 



LEGEND 
— <^ Watershed drainage divide 
i2» Monitored ground water supply 
H ** Stream monitoring location 
Y/A Mined longwall panel 
6-1-86 Date ot mining 
Stream 

3,000 

I i _J 



Scale, ft 
FIGURE 2.— Location of monitored ground water supplies and streams at mine site Y. 



75 



utilized for monitoring flow at stream H 
of mine Y owing to its relatively large 
size. 

The locations of the sampled water sup- 
plies and stream monitoring points were 
plotted on the appropriate 7.5-min topo- 
graphic map along with the mine outline 
and type of mining method used. Water- 
shed boundaries were then determined and 
plotted on this map for each stream moni- 
toring point. The watershed boundaries 
were constructed by first drawing two 
lines uphill in either direction from the 
stream monitoring point perpendicular to 
the elevation contours until the nearest 
watershed ridge top was intersected, and 
then tracing the topographic drainage 
divide along the ridge tops. The area 
of each watershed was finally determined 



by utilizing a dot matrix counting 
technique. 

The sampled wells, springs, and stream 
points were then plotted on various fig- 
ures to determine the horizontal and ver- 
tical spacings between these features and 
the mine. The stratigraphic position of 
each water supply and stream point was 
also determined. Finally, horizontal and 
vertical spacings were measured from the 
sampled water supplies to both the near- 
est stream and the regional base level, 
and the topographic position of these 
supplies was also determined. This in- 
formation was then used to determine the 
extent of strata dewatering and to quan- 
tify the relationships of certain hydro- 
geologic parameters with dewatering. 



RESULTS FOR MINE Y 



PREMINING HYDROLOGIC BALANCE 

Hydrologic conditions existing prior to 
mining at the mine Y site were determined 
by examining nearby local streams and wa- 
ter supplies that were beyond the influ- 
ence of mining and up to 2,000 ft away 
from the nearest mine. The shallow water 
table for hilltops was typically encoun- 
tered at a depth of 10 ft, and the shal- 
low water table under the stream valleys 
was encountered at depths of 10 ft and 
less. The depth to static level in deep- 
er wells averaged 25 ft in the stream 
valleys and 100 ft on the hilltops. The 
most dependable water wells were devel- 
oped into the Jollytown Sandstone Unit in 
the western portion of the mine. In the 
eastern portion of the mine the Waynes- 
burg Sandstone was the most prolific 
aquifer. These units yielded the most 
dependable water supplies, especially 
when the water wells developed in them 
were located within 100 ft vertically of 
a stream valley. In two cases such wells 



were artesian, and even nonartesian 
valley wells often provide enough water 
for four or five families each. Devel- 
oped springs and water wells that are 
located on hill tops or upper hill slopes 
appear to have more variable, less de- 
pendable yields or water levels affected 
by seasonal changes in precipitation and 
evapo transpiration. 

Based on the analysis of control 
streams in watersheds that were not un- 
dermined, the bulk of the streams can be 
classified as small perennial streams. 
During dry periods stream discharges were 
composed of just ground water runoff 
which usually sustained some measurable 
streamflow. Such streamflow (or base- 
flow) was proportional to watershed area. 
The normal median streamflow calculated 
relative to watershed area from baseflow 
measurements was 0.0645 gal/(min*acre) on 
June 6, 1986, and 0.0136 gal/(min*acre) 
on July 24, 1986, during a much drier 
period than in June. 



76 



Stream G 



Streamy 



H6 



Unmined coal 




6-85 



9-84 



9-83 



LEGEND 
Watershed drainage divide 
i2» Monitored ground water supply 
H ** Stream monitoring location 
V7A Mined longwall panel 
P2 Longwall panel 
7- 24-86 Date ot mining 

Stream 

2,000 



Scale, ft 



FIGURE 3.— Location of monitored ground water supplies and streams overlying or in close proximity to longwall panels at 
mine site Y. 



77 



MINING IMPACTS ON HYDROLOGY 
Wells and Springs 

Water supplies overlying or in near 
proximity to room-and-pillar or mine en- 
try sections have had no reported nega- 
tive impacts associated with this type of 
mining. This situation appears to be in 
response to the relatively thick mine 
overburden of 490 ft or more between 
these water supplies and mine Y. 

Analysis of the water supplies devel- 
oped in the subsided strata overlying 
longwall panels (as located in figure 3) 
indicated that 8 of 11 domestic water 
supplies that were monitored both be- 
fore and after mining were partly to 
completely dewatered (fig. 4). The maxi- 
mum amount of dewatering appears to 
have been more extensive near longwall 
panel centers, such as zone 3 of figure 
4, judging by the lowest position of com- 
pletely dewatered supplies. Dewatering 
appears to have been limited to the 
strata located at least 655 to 700 ft 
above the base of the Pittsburgh Coal. 
An examination of the ratio of panel 
width to mine overburden thickness (fig. 
that dewatering at mine Y 
a range of ratio values of 
This range of values cor- 
a minimum mine overburden 
thickness of 655 ft (as shown in figure 
4), above which dewatering was reported 
to have occurred. 

Water supplies located adjacent to but 
not above longwall panels were also exam- 
ined for dewatering trends. Analysis of 
these supplies determined that dewater- 
ing zones were present, as defined by an 
angle of influence (fig. 6). The angle 
of influence is defined as the angle be- 
tween a vertical line projected upward 
from the edge (rib or end) of a longwall 
panel and a line projected to the fur- 
thest point of dewatering effects from 
the longwall panel (6). Eleven of 13 



5) indicated 
occurred over 
0.75 to 1.00. 
responded to 



900 



Q 
UJUI 

via: 

DOtl 



yo-8C 



KEY 
Monitored ground water supplies undermined 
by longwall panel (a) 

a No dewatering 

a Partial dewatering 

■ Complete dewatering 

a, i Partial recovery 

Dewatering and recovery unknown 

Drilled or dug wells (with varying degrees of 
development, recovery, and / or dewatering) 

b Ground water supply developed post mining 

Monitored water supply 



i 



etc 



700- 



C53 

ZCo 

en & 
^ | - 

03 |_ 



600- 



500 



400 



930 H '29 


|H 


1 


I 


i3 


31 










c 


■ 




h ' 


Dewatering ~ 
zone 




I8-QI9 










I 


__*»_ 


15 

■ 




"i — 


a 
oie 


58 07 
oo 
10 




1 


No 

> dewatering 

zone 


D5 


A 






. 


Zone 1 


Zone 2 

I 


Zone 3 


i i 



100 200 300 400 500 

HORIZONTAL DISTANCE OF MONITORED 

GROUND WATER SUPPLY INSIDE FROM NEAREST 

EDGE OF LONGWALL PANEL, ft 

FIGURE 4.— Vertical extent of dewatering for wells and 
springs over longwall panels at mine site Y. 






i r 



KEY 
Monitored ground water 
supplies at mine Y 







W/ 


1^3 Complete dewatering 




y///, 


mm Partial dewatering 




//// 

m 


m 


I iNo dewaterinc 


i 


m 


W/< 


:::::i 


i 







0.75 0.85 0.95 1.05 1.15 1.25 

RATIO OF PANEL WIDTH TO OVERBURDEN THICKNESS 

FIGURE 5.— Dewatering as a function of the ratio of 
longwall panel width to mine overburden thickness for 
monitored ground water suplies at mine site Y. 



78 



KEY 
Monitored ground water supplies undermined by room-and-pillar 
section (a), or not undermined (A) 

Drilled or dug wells (with varying 
degrees of development, recovery, 
,etc and /or dewatering) 

o Ground water supply developed 
post mining 

28 Monitored water supply 



A No dewatering 

A, A Partial dewatering 

A, A Complete dewatering 

ik, A Partial recovery 

M, Jik Complete recovery 



A^,-A^ Dewatering and recovery 
unknown 




*s^l2 I Dewatering 
zone 



dewatering nr 
zone 

A5 



200 400 600 800 1,000 1,200 1,400 
HORIZONTAL DISTANCE FROM NEAREST LONGWALL PANEL, ft 

FIGURE 6.— Extent of dewatering for water supplies adjacent to longwall panels at mine site Y. 



domestic water supplies that were moni- 
tored both before and after mining and 
located within the maximum 42° angle of 
influence zone were reported to have been 
partly to completely dewatered by the 
longwall mining (fig. 6). This dewatered 
zone adjacent to longwall panels appears 
to be limited to strata located at least 
700 to 720 ft above the base of the 
Pittsburgh Seam, judging by affected wa- 
ter supplies that are located in or most- 
ly in such strata. 



Partial to complete dewatering of water 
supplies was found to usually extend no 
lower stratigraphically than the base of 
the Jollytown Sandstone (fig. 7). Strat- 
igraphic cross section A-A'-A" extends 
the length of the center of panel 2 as 
indicated in figure 2. The Jollytown 
Sandstone is situated at or just above 
the hypothesized regional base level 
that exists at the position of the major 
stream (stream H) for this portion of 
Mine Y. The extent of strata dewatering 



79 



KEY 
Location of ground water supplies with respect to stratigraphic cross section A-A-A" 
monitored ground water supplies undermined by longwall panel (p), or by room-and- 
pillar section (a) 
a, a No dewatering 
A i a Partial dewatering 
A - ■ Complete dewatering 
A, A, a, k Partial recovery 

-A,- Dewatering and recovery unknown 



o 

25 



Drilled or dug wells (with varying degrees of 
development, recovery, and/or dewatering ) 
Ground water supply developed post mining 
Monitored water supply 



UJ 
> 

UJ 

< 

L±J 
if) 



< 
UJ 



LU 
> 

o 

< 



< 
> 

UJ 

_1 
UJ 










c 


a> 


o 


c 




a> 


D 


a> 


h 




t- 


(J3 


o 




u. 






c 


c 


o 


o 






c 


£ 



1,000 1,500 2,000 2,500 \ 4,500 5,000 5,500| 

FIGURE 7.— Location of ground water supplies with respect to stratigraphic cross section AAA" at mine site Y. 



with respect to regional base level is 
shown in figure 8 for supplies located 
reasonably close to stream H. Strata de- 
watering was found to extend down to a 
level located about 50 ft above the re- 
gional base level defined by stream H. 
Water supplies developed mostly below 
this 50-ft level appeared to be unaf- 
fected by the mining. Figure 9 shows 
that water supplies located mostly above 
the level of nearby perennial streams 
(not counting stream H) were dewatered 
within the 42° angle of influence zone 
adjacent to longwall panels. Those water 
supplies developed mostly below nearby 



streams in elevation appear to have 
been unaffected by the mining. Perennial 
streams undermined by longwall panels, as 
shown in figure 10, seem to be associated 
with more strata dewatering than such 
streams not over panels (fig. 9). Dewa- 
tered supplies extend to about 20 ft or 
more below the elevation of nearby panel 
streams, with the maximum dewatering ex- 
tending to 60 ft below stream elevation 
at the panel center position, as indi- 
cated by the water level positions mea- 
sured by Moebs and Barton (_1_) and Walker, 
Green, and Trevits (7). 



80 



KEY 
Monitored ground water supplies undermined 
by longwall panel (□), or by room-and-pillar 
section (a) 

a, a No dewatering 
A, ■ Complete dewatering 
4 Partial recovery 
-A Dewatering and recovery unknown 
o-a, A— A Drilled or dug wells 

o Ground water supply developed post mining 

58 Monitored water supply 

900 



800 



.r 700 



< 




Ll) 




oc 


600 


\- 




<f) 




2 




o 




an 


500 


u. 




UJ 




o 




z 




£ 


400 


co 




a 




_i 




< 


300 


Z 




o 




N 




or 
o 


200 


X 





100 - 



No dewatering 
zone 



Dewatering 
zone 



-A35 



ill 



— D6 

-a 7 



H5 



9oo58 
l0 ° 057 



OI6 
-*8 I 




-200 



I 



-100 







100 



200 300 



VERTICAL DISTANCE ABOVE STREAM //, ft 

FIGURE 8.— Extent of dewatering for ground water supplies 
with respect to their vertical position above stream H. These 
supplies are nearer to stream H than to any other stream at 
mine site Y. 



KEY 
Monitored ground water supplies located out- 
side longwall panel 

□ No dewatering 

a Partial dewatering 

■ Complete dewatering 

A, 4 Partial recovery 

-a,-o Dewatering and recovery unknown 

o — o,d— a Drilled or dug wells (with varying degrees of 
-n— q,b— a development, recovery, and /or dewatering) 



o 
24 



CO 
Ui 

or 
< 

UJ 



900 



800 



700 



UJ *- 

R* 

uj°- 

O-l 
Z-l 

?% 

CO CD 



600 



500 



400 



o 

N 

or 
o 

X 



300 



200 



00 



Ground water supply developed post mining 
Monitored water supply 

— I — 



T 



T 



T 



No dewatering Dewatering 
zone zone 



29-a- 



! 23^ 



23 

Q4 i 



i26 



24o o i 22t 25 



3 4-4 



I4o4 



24- 



l7o 



I 



I 



-D59 



_L 



-200 -100 100 200 300 

VERTICAL DISTANCE ABOVE NEAREST 
PERENNIAL STREAM, ft 

FIGURE 9.— Extent of dewatering for ground water supplies 
not over longwall panels with respect to their vertical position 
above the nearest perennial stream at mine site Y. 



81 



KEY 

Monitored ground water supplies undermined 
by longwall panel 
□ No dewatering 
a Partial dewatering 
■ Complete dewatering 
a, i Partial recovery 
-c,-B Dewatering and recovery unknown 

0—0,0—0 Drilled or dug wells (with varying degrees of 
a— a, etc. development, recovery, and /or dewatering ) 

o Ground water supply developed post mining 

19 Monitored water supply 



CO 

LJ 

< 

UJ 

z ^_ 
UJ* 1 

Sj 

CO UJ 

z z 
-<t 

UJ 0- 

d-z 



o 

N 

rr 
o 

X 



V 


Panel 
edge 


No dewaterinc 
zone 


1 Dewatering 
zone 








3i^ 


a 








5D- 




D 

-■18 






100 








19-u ° 


*3I 




200 


Panel 




/ 
/ 






- 




center 












{ 1 




\±J— 








inn 


H a 1 


1 



-300 -200 



100 



1 00 200 



300 



VERTICAL DISTANCE ABOVE NEAREST 
PERENNIAL STREAM, ft 



FIGURE 10.— Extent of dewatering for ground water sup- 
plies above longwall panels with respect to their vertical posi- 
tion above the nearest perennial stream at mine site Y. 



The previous results and supply dewa- 
tering trends show that zones of longwall 
mine subsidence have been associated with 
extensive strata dewatering extending 
down to or below the levels of nearby 
streams but not quite to regional base 
level as defined by stream H. This prob- 
ably indicates that increased vertical 
hydraulic conductivity associated with 
subsidence fractures has allowed an ex- 
tensive general lowering of ground water 
levels to occur under hills and ridges. 



The ultimate fate of this drained ground 
water will be addressed later. 

Streams 

Measured streamflow data were next 
examined for any dewatering trends asso- 
ciated with the underground mine. The 
positions of the streams and their 
monitoring points are shown in figures 2 
and 3, and the stream discharge on June 
6, 1986, as a function of watershed area 
for the subwatersheds is plotted in fig- 
ure 11. The stream A watershed was 
divided into three subwatersheds defined 
by stream B downstream to point Bl, 
stream C downstream to point CI, and 
stream A downstream of points Bl and CI. 
Subwatershed Bl was monitored at three 
different locations on June 6, 1986. At 
monitoring point B3, streamflow appeared 
to be normal. The stream discharge to 
watershed area ratio of 0.062 gal/ 
(min'acre) was within the normal range of 
median streamflow to watershed area ra- 
tio, 0.050 to 0.088 gal/(min»acre), as 
shown in figure 11 for the streams not 
impacted by mining. In contrast, subwa- 
tershed B2-B3 appeared to have lost wa- 
ter, whereas stream discharge at moni- 
toring point B3 was 4.75 gal/min, stream 
discharge at the downstream monitoring 
point B2 was only 2.00 gal/min. The de- 
crease in stream discharge from B3 to B2 
was reflected in the stream discharge in- 
crease to watershed area ratio of -0.688 
gal(min*acre) , which was about 0.753 gal/ 
(min # acre) less than normal (fig. 12). 
At monitoring point Bl, the stream bed 
was completely dry. The calculated 
stream discharge increase to watershed 
area ratio of stream B1-B2 was -0.667, or 
about 0.733 gal/(min*acre) less than nor- 
mal (fig. 12). Subwatershed CI was com- 
pletely dry on both June 6 and July 24, 
1986, and no measurable streamflow was 
apparent at monitoring points C2 and C3 
on July 24 as well. 



82 



14 



12 



c 

E 


10 


^ 




o 




CT 




ft 




LlJ 


B 


o 




cc 




< 




X 




(J 




en 


6 


Q 




2 




< 




UJ 

ce 


4 


\- 




co 





uj 2 

UJ 
CC 

^ 



-2S- 



-4 



T 



Normal streamflow zone 
with no dewatering 



Increased streamflow 
zone 




^Ao 
















Decreased streamflow 
zone with dewatering 



A3-BI-CI 



'CI 



BI-B2 



JB2-B3 
IGI-G4 



1 



50 



100 150 200 

WATERSHED AREA, acres 



250 



300 



FIGURE 11.— Stream discharge Increase versus sub-watershed drainage area for baseflow of June 6, 1986, for control and min- 
ing affected streams at mine site Y. 



Subwatershed A1-B1-C1 was measured at 
three points on July 24, 1986. Monitor- 
ing point Al was located 10 ft upstream 
from its junction with stream H. Moni- 
toring point A2 was located 44 ft up- 
stream from Al , and monitoring point A3 
was located 50 ft upstream from point A2. 
A portion of the A1-B1-C1 watershed, A3- 
Bl-Cl, was completely dry when monitored 
June 6 and June 24, 1986. Directly down- 
stream from point A3, the Jolly town Sand- 
stone crops out in the streambed. Be- 
teen this outcrop and monitoring point 



A2, water was pooled with no measurable 
streamflow on July 24. In contrast, this 
stream section exhibited a high stream- 
flow to watershed area ratio of 0.714 
gal/(min*acre) on June 6, with flow 
starting farther upstream immediately 
downstream from point A3 during the less 
dry conditions. Stream segment A1-A2 had 
a measurable streamflow of 5.40 gal/min 
on July 24, with a calculated stream dis- 
charge to watershed area ratio of 2.08 
gal/(min*acre). Streamflow at point Al 
was not measured on June 6. 



83 



>- 




00 






^_ 


O 


u 


UJ 


Q. 


2 


^ 


1 


UJ 


(Z 


7* 


Id 

o 


2 




_J 
_J 


Q 


< 


UJ 


$ 


X 


(.9 


tn 


F" 


Lt 


o 


UJ 


_l 


I- 




< 




£ 





100 



80 



60 



KEY 
B2m Stream sampling point 



\ 



40 



20 







BI-B2 
>B2-B3 



_Decreased streamflow_ 
zone with dewatering 




fT 



i i 

Normal streamf low zone 
with no dewatering 



Increased 



streamf low zone 



• A2-A3 



EI-E2,F2, Gil, FI-F2 
Dl, D2 I 



•0.80 -0.60 -0.40 -0.20 



0.20 



0.40 0.60 



RATIO OF ACTUAL MINUS MEDIAN NORMAL STREAM 
DISCHARGE TO WATERSHED AREA, gal/(min- acre) 

FIGURE 12.— Ratio of actual minus median normal stream discharge to watershed area versus percentage of watershed under- 
mined by longwall panels at mine Y. 



Stream G (fig. 3) was monitored on June 
6 at 4 sites and on July 24 at 10 sites 
to test for a possible angle of influence 
effect on this watershed from panel 3; 
see figures 12 and 13 for plotted stream 
data for June 6, and July 24, respective- 
ly. Figure 13 shows stream segment G1-G2 
near stream H to have been completely dry 
on July 24. Upstream, segment G2-G3 had 
pooled water with no measurable flow. 
Between stations G3 and G9 the stream ap- 
peared to have been steadily losing water 
as it flowed downstream toward stream H. 
As shown in figure 13, there was an 



apparent angle of influence of 31°, as 
measured from the panel end to point G9 
where streamflow had peaked for stream G. 
Stream G lost water in this 31° dewater- 
ing zone as well as in the 21° zone lo- 
cated inward from the panel end (fig. 
13). Upstream from station G9 and beyond 
the influence of panel 3 there appeared 
to be normal streamflow. 

Stream H, the major stream in the mined 
study area, was monitored on July 24 at 
two locations. Monitoring station HI was 
located downstream from the area of long- 
wall mining, and monitoring station H6 



84 



X 




o 




en 




ZD 




00 




CO 




\- 




\- 


»♦— 


n 






CO 


UJ 


r- 


X 


Z 


H 


d 


U. 


0_ 


o 


o 


UJ 


z 


CO 


_l 


< 


CL 


00 


s 


-z. 


< 

CO 


UJ 




UJ 


^ 


^ 


< 


h- 


111 


UJ 


a: 


00 


\- 




co 


(D 




CO 


o 


III 


z 


z 


< 


o 


Q 
UJ 


X 


CO 


r- 


_l 




< 


2 

UJ 


o 
o 


Q 




LT 




3 




00 




C£ 




UJ 




> 




o 





800 



700 



? 600 



500- 



400- 



300 



200- 



100- 







Pooled water 



1 KEY ' 
65 a Stream sampling point 




610 



" Angle for 
complete 
stream loss Vr 



Angle for partial 
stream loss- 



Longwall panel 



-300 -200 




01 



uj2 

C3|_ 

<cr 

CO UJ 

So: 
< 



235 6789 10 
WATERSHED 6 STREAM - 
SAMPLIN6 POINTS 



-100 



100 



200 



300 



400 



500 600 



DISTANCE OF STREAM SAMPLING POINTS FROM EDGE OF PANEL, ft 

FIGURE 13.— Discharge of stream G as a function of location with respect to the end of panel 3 at mine Y. 



was located upstream from the area of 
longwall mining; see figure 2 or 3. 
Stream segment H1-H6 was calculated to be 
gaining 44.0 gal/min. To determine any 
possible impacts of mining on stream seg- 
ment H1-H6, its flow rate was compared to 
that of stream El, a control stream that 
was not impacted by mining, as shown by 
figures 11 and 12. Control stream El was 
selected for comparison to stream segment 
H1-H6 over control stream G10, which was 
also not affected by mining, because 
stream El is closer in size of drainage 



area to segment H1-H6 and because it has 
a more similar geologic setting to seg- 
ment H1-H6. The stream discharge to 
watershed area ratio at monitoring point 
El, was 0.024 gal/(min*acre) on July 24. 
This value was then multiplied by the wa- 
tershed area of subwatershed H1-H6 to 
yield an expected normal stream discharge 
of 26.3 gal/min. The difference between 
the measured stream gain of 44.0 gal/min 
and the 26.3 gal/min represented an ex- 
cessive streamflow gain of 17.7 gal/min 
between monitoring points HI and H6. 



85 



This excessive streamflow indicates that 
subwatershed H1-H6 had a 67 pet greater 
streamflow than expected compared to the 
"normal" El watershed. Although natural 
geologic differences may account for some 
of this difference in streamflow related 
to ground water runoff contributions, 
most of this extra flow was probably due 
to the influence of mine Y. 

See figure 14 for the projected strati- 
graphic locations of these monitored 
stream points for the mine Y area. This 
figure is similar to figure 7, discussed 
earlier. Although stream station El 
could not be plotted on figure 14, it 
is located stratigraphically below the 
Jollytown Sandstone. 

Fate of Lost Water Over the Mine 

The schematic vertical cross section 
(fig. 15), which is based on dewatering 
evidence collected from water wells, 
springs, and streams at mine Y, illus- 
trates a generalized lowering of the 



potentioraetric surface in response to the 
longwall mining. The dewatering data in- 
dicate that the water lost from the sur- 
face to near-surface zones was recharging 
down to the Jollytown Sandstone or adja- 
cent strata located immediately above the 
regional base level. This probably oc- 
curred because of tensional subsidence 
fractures in the hills which had signifi- 
cantly increased the vertical permeabil- 
ity and facilitated the rapid downward 
drainage of water. Although no obvious 
fractures were observed in the soil zone 
above mine Y, mine subsidence theory 
nonetheless predicts their presence in 
bedrock of an upper fractured zone or 
zone of surface rock fracturing (J3). 
This lost water could have then possibly 
flowed downdip along the Jollytown Sand- 
stone to discharge into major streams in 
the local area. 

There does not appear to be any evi- 
dence that these lost waters recharged to 
the underground mine. Evidence against 
such deep mine recharge is the probable 





1,300 


H- 




_l 




UJ 




> 

UJ 


1,200 


_J 




< 




UJ 




(0 




z 


1,100 


< 




UJ 




2 




UJ 


1,000 


o 

CD 




< 




z 
o 


900 


r- 




< 




> 




UJ 





Ul 



800 




500 1,000 1,500 2,000 2,500 3,000 3,500 

DISTANCE, ft 



FIGURE 14.— Location of selected stream monitoring stations with respect to stratigraphic cross section A-A' at mine Y. 



86 



Premining 

potentiometric 

surface 



Maximum mining- 
impacted 
potentiometric 




FIGURE 15.— Schematic vertical cross section showing generalized dewatering effects at mine Y. 



presence of an aquiclude zone located 
below the regional base level (fig. 15). 
That zone is predicted by mine subsidence 
theory, since the minimum mine overburden 
thickness (500 ft) far exceeds the proba- 
ble vertical extent of the lower frac- 
tured and caved zone [30 to 60 times the 
mined coal seam thickness (8)]. Also, 
the greater lateral compressive stress 
and elastic rebound of subsided strata 
within the aquiclude zone should keep 
fracture permeability relatively low 
there (8^. Lack of significant vertical 
movement of water through this zone is 
indicated by several nondewatered wells 
within it and the formation of five new 
springs just above it. Additional evi- 
dence against deep mine recharge is that 
company personnel at mine Y reported that 
this mine received only about 0.01 gal/ 
(min*acre) of recharge. 



Stream H was investigated as a likely 
major stream to be receiving lost water 
runoff. The apparent excessive increase 
in streamflow for stream H must be ac- 
counted for by the inclusion of an addi- 
tional water source that is not normally 
a part of the stream's hydrologic budget. 
Otherwise, stream H would have had a nor- 
mal stream discharge to watershed area 
ratio, which would have been typical for 
an unmined stream in the area. The most 
probable additional water source was ex- 
tra water saved from evapotranspiration 
due to the effect of mining. 

To determine the possibility of an 
evapotranspiration saving accounting for 
the increased streamflow at stream H, it 
was first necessary to estimate the evap- 
otranspiration rate (ET) for the study 
area. The U.S. Geological Survey (Pitts- 
burgh office) has reported average 



precipitation to be 39.4 in/yr for the 
period 1941-80 based on a rain gauge at 
Waynesburg near mine Y (_5)« They also 
estimated that ET averaged 61 pet of to- 
tal precipitation for this same period 
for a portion of the watershed of South 
Fork Tenmile Creek that includes mine Y 
(_5_). Based on these data, the ET should 
have averaged 24.0 in/yr for the 1941-80 
period at mine Y. This is equivalent to 
6.52 x 10 +5 gal/(acre*yr). That ET value 
was then multiplied by the 535 acres of 
longwall mine area (including the zone of 
the 42° angle of dewatering influence) 
that could have been contributing to 
ground water runoff entering stream H. 
This produced a yield of 3.49 x 10 +8 gal/ 
yr, or 663 gal/min, of average ET from 
the longwall mine area. The ratio of ex- 
cess streamflow in the measured stream H 
segment (17.7 gal/min) to the average ET 
over the mine is 2.7 pet. Therefore that 
excess streamflow could be accounted for 
by a 2.7-pct saving in ET over the mine. 
This water saving could have been real- 
ized following subsidence by a diver- 
sion of shallow ground and stream waters 
along subsidence fractures to deeper lev- 
els closer to the regional base level, 
where the water would be less suscepti- 
ble to evaporation or transpiration by 
plants. The water that was thereby 
spared from ET would have been added to 
the ground water runoff portion of 
streamflow for stream H. 

WATER RECOVERY FOLLOWING MINING 

Some recovery of water levels occurred 
for domestic water supplies and streams 
following mining at mine Y. Of the two 
accessible water supplies that had been 
partly dewatered over longwall panels, 
both showed partial but not complete re- 
coveries. Of the four accessible sup- 
plies that were completely dewatered over 
longwall panels, only two had a partial 
recovery and none had a complete recov- 
ery. Of nine affected water supplies for 
which recovery data were available and 
which were not undermined by a longwall 
panel but were located within the zone of 
the 42° angle of influence, eight had a 



partial recovery and one had a complete 
recovery. Water supplies that showed 
partial to complete recoveries did so 
within 1 to 3 yr after dewatering oc- 
curred and were typically located close 
to or below nearby perennial streams or 
hydrologic base level. Supplies located 
over panel centers showed less overall 
recovery compared to supplies located 
near panel edges or adjacent to panels. 
Only one supply (well 2) has exhibited 
complete recovery so far. 

While some old affected springs showed 
partial recovery, several new springs ap- 
peared or were developed at lower eleva- 
tions following mining. Two new seasonal 
springs were developed over panel 2; one 
new perennial spring was developed over 
panel 3 and is being used to supply five 
families; one new seasonal spring was de- 
veloped over panel 4; and one new spring 
was developed over panel 5. All of these 
springs appeared immediately to 1 yr 
after undermining occurred. The new 
springs appearing above panels 2 and 3 
are located stratigraphically at the top 
of the Jollytown Sandstone, whereas the 
new springs above panels 4 and 5 are 
situated approximately 20 ft above the 
stratigraphic level of the Jollytown 
Sandstone (fig. 14). In addition, a do- 
mestic well that was not dewatered over 
the mine had a reported increase in yield 
following mining; this well penetrates to 
below the Jollytown Sandstone. These 
enhanced water supplies again indicate 
that overlying aquifer waters were di- 
verted downward and that partial recovery 
of ground water often occurred close to 
the local or regional base (stream) level 
within 1 yr of mining. 

A comparative analysis of streamflow to 
watershed area ratios for June showed in- 
directly that panel 1 had a complete re- 
covery and panels 2 and 3 had partial re- 
coveries since longwall mining occurred. 
The portion of panel 1 that undermined 
subwatershed B3 was between 3.0 and 3.5 
yr old, whereas the portion of panel 2 
that undermined subwatershed B2 to B3 was 
between 2.0 and 2.5 yr old. Subwatershed 
B2 to B3 was calculated to be losing 
0.811 gal/(min*acre of longwall panel) to 



88 



the subsurface. In contrast, the portion 
of panel 3 that undermined the stream G 
watershed was between 1 and 2 yr old, and 
subwatershed G4 to Gil was calculated to 
be losing 2.08 gal/(min*acre of long- 
wall panel) to the subsurface. No cal- 
culations of maximum streamflow loss, 
such as those above, were possible for 



panels 4 and 5 owing to the dry streams 
encountered over these panels. The obvi- 
ous trend from these data is that streams 
show progressive recovery with time 
between about 1 and 3 yr of longwall min- 
ing and recover their normal premining 
flow by about 3 yr after mining. 



CONCLUSIONS FOR MINE Y 



1. Roora-and-pillar mining (with no 
subsidence) has not affected water sup- 
plies or streams. 

2. Above longwall panels, 8 of 11 wa- 
ter supplies appeared to have been partly 
to completely dewatered. Dewatering oc- 
curred in a zone about 655 ft and higher 
above the mined Pittsburgh Coal Seam, but 
not closer to this seam. Dewatering was 
greatest above the center of the longwall 
panels. All supply dewatering over long- 
wall panels occurred within a range of 
0.75 to 1.00 for the ratio of longwall 
panel width to mine overburden thickness. 

3. Off of and away from the longwall 
panels, 11 of 13 water supplies were 
partly to completely dewatered within a 
maximum angle of dewatering influence of 
42° from the nearby panel ribs or ends. 
Dewatering occurred in a zone of about 
720 ft and higher above the mined Pitts- 
burgh Coal Seam, but not within 720 ft of 
this seam. 

4. Dewatering did not occur below the 
regional base level, as represented by 
the major stream over the mine. Water 
supplies developed within 800 ft lateral- 
ly of the major stream and no higher than 
50 ft above it were not dewatered, but 
such supplies over 50 ft above it were 
partly to totally dewatered. 

5. Water supplies developed at or be- 
low the elevation of the nearest peren- 
nial stream and outside of panels but in- 
side the 42° angle of influence zone did 
not appear to be dewatered. However, 
such supplies developed mostly above the 
level of the nearest perennial stream 
were at least partly dewatered. Most wa- 
ter wells developed more than 20 ft below 
the elevation of the nearest perennial 
stream and undermined by longwall panels 



were not dewatered, but such wells lo- 
cated above this level were partly to 
totally dewatered. Above the center of 
one longwall panel, a well was dewatered 
to a depth of 60 ft below the elevation 
of the nearest perennial stream but not 
below regional base level. 

6. Streams located above regional base 
level and undermined by longwall panels 
less than 2.5 yr old were partly to com- 
pletely dewatered during baseflow condi- 
tions. An angle of dewatering influence 
of 31° existed outside the end of such a 
panel for one measured stream, with par- 
tial stream dewatering having occurred 
within this area adjacent to the panel. 
Streams located above regional base level 
and also above panels at least 3 yr old 
had normal flows. 

7. Water from lost ground water sup- 
plies and streams did not penetrate to 
the mine, but instead migrated downward 
through probable subsidence fractures to 
near regional base level, where it mi- 
grated laterally through the Jollytown 
Sandstone to finally discharge to the 
largest area stream over the mine. Re- 
ported ground water recharge to the deep 
mine was only 0.0.1 gal/(min*acre). The 
major area stream flowing over the mine 
received about 17.7 gal/min more than the 
expected normal runoff contribution on 
July 24, 1986. This extra slug of runoff 
during dry baseflow conditions was most 
likely contributed by a potential evapo- 
transpiration reduction of about 2.7 pet 
over the mine that resulted from reduced 
stream and ground water levels at higher 
elevations. 

8. Several streams and water supplies 
showed partial to total recovery follow- 
ing mining effects. Streams appeared to 



89 



have had complete recovery within 3 yr 
after longwall mining occurred. Of the 
accessible ground water supplies over 
longwall panels, all partly dewatered 
supplies had partial recovery, but only 
one-half of the completely dewatered sup- 
plies had a partial recovery, with no 
complete recovery observed. Of the af- 
fected accessible supplies adjacent to 
longwall panels, all showed at least par- 
tial recovery. Several new springs also 
formed close to stream level within 1 yr 



of nearby mining. Any noted recovery for 
affected wells and springs occurred with- 
in 1 to 3 yr of the initial mining ef- 
fects, with more extensive and rapid re- 
covery for supplies near stream level or 
not over panel centers. Only one af- 
fected water supply had shown complete 
recovery as of February 1986. Affected 
streams exhibited more complete recovery 
than did wells and springs following 
mining. 



REFERENCES 



1. Moebs, N. M. , and T. M. Barton. 
Short-Term Effects of Longwall Mining on 
Shallow Water Sources. Paper in Mine 
Subsidence Control. Proceedings of Tech- 
nology Transfer Seminar, Pittsburgh, PA, 
September 19, 1985, comp. by Staff, Bu- 
reau of Mines. BuMines IC 9042, 1985, 
pp. 13-24. 

2. Merritt, P. C. U.S. Sees Marked 
Growth in New Longwall in 1985 (1985 
Longwall Census). Coal Age, v. 90, No. 
8, 1985, pp. 47-65. 

3. Tieman, G. E. Study of Dewatering 
Effect at Two Underground Longwall Mine 
Sites in the Pittsburgh Coal Seam of the 
Northern Appalachian Coal Field. Unpub- 
lished M.S. thesis in Geology. WV Univ., 
Morgantown, WV, 1986, 147 pp. 

4. Stoner, J. D. Probable Hydrologic 
Effect of Subsurface Mining. Ground Wa- 
ter Monitoring Review, v. 3, No. 1, 1983, 
pp. 128-137. 

5. Stoner, J. D. , D. R. Williams, 
T. F. Buckwalter, J. K. Felbinger, and 
K. L. Pattison. Hydrogeology , Water 



Resources, and the Hydrologic Effects of 
Coal Mining, Greene County, Pennsylvania. 
U.S. Geol. Surv. , 4th Ser. , Water Resour. 
Rep., PA Geol. Surv., (in press, 1987). 

6. Cifelli, R. C. , and H. W. Rauch. 
Dewatering Effects From Selected Under- 
ground Coal Mines in North-central West 
Virginia. Paper in Proceedings of Second 
Workshop on Surface Subsidence Due to 
Underground Mining. WV Univ. , Morgan- 
town, WV, 1986, pp. 249-263. 

7. Walker, J. S., J. B. Greene, and 
M. A. Trevits. A Case Study of Water 
Level Fluctuations Over a Series of Long- 
wall Panels in the Northern Appalachian 
Coal Region. Paper in Proceedings of 
Second Workshop on Surface Subsidence Due 
to Underground Mining. WV Univ. , Morgan- 
town, WV, 1986, pp. 264-269. 

8. Coe, C. J., and S. M. Stowe. Eval- 
uating the Impact of Longwall Coal Mining 
on the Hydrologic Balance. Paper in Pro- 
ceedings of Conference on the Impact of 
Mining on Ground Water. National Water 
Well Association, 1984, pp. 348-359. 



1187 234 



US GOVERNMENT PRINTING OFFICE 1987 605 01 7 6<X)49 



INT.-BU.OF MINES,PGH.,PA. 28491 





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